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SECOND OOPY, 
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NOTES 



ON 



Lead and Copper Smelting 



AND 



COPPER CONVERTING-. 



BY 



HIKAM W. HIXO^^ 



Late Superintendent of the Blast Furnace and Converter Department, 
Anaconda, Montana. 



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NEW YORK AND LONDON: 
THE SCIENTIFIC PUBLISHING CO., 

1897. 






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38163 



Copyrighted 1897, 

BY 

THE SCIENTIFIC PUBLISHING COMPANY, 




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PREFACE. 



This book is precisely what is indicated by its title — a series 
of notes on the practical work in lead and copper smelting, 
includino; the converting of copper matte. It is by no means a 
treatise or an attempt at a treatise. No effort has been made 
to trace in order all the steps in beneficiating ores by smelt- 
ing from the crude material to the marketable product. This 
has already been done ably by other writers. It has seemed, 
however, that the experience gained in the everyday operation 
of three large works, extending over a period of ten years, 
might be useful to others who are engaged in similar work. 
Progress in any art is helped by an interchange of ideas. 
Hence these notes are offered. 

Acknowledgement is due Mr. Wm. Braden for the reproduc- 
tion of drawings of the settlers used at the Arkansas Valley 
Smelting Works, Leadville, Colo., from his paper in the 
Transactions of the American Institute of Mining Engineers^ 
Vol. XXVI ; and to L. S. Austin for the illustrations of the 
matte-pots and slag-trucks employed at the Omaha & Grant 
Smelting Works, Denver, Colo., which are taken from his 
paper on the separation and disposal of slag in The Engineer- 
ing and Mining Journal of November 23, 1895 ; also to Julius 
A. Dyblie and John Bendixen for valuable services in prepar- 
ation of plans and drawings. 



TABLE OF CONTENTS. 



Chapter I.— Copper Matte Smelting. 

Percentage of copper required — Alteration of lead furnaces 
for copper matte smelting at Leadville, Colo.— Effect of 
fine ore in the charge — Kind of material smelted at Lead- 
ville — Composition of slag — Proportion of matte-fall 
necessary in copper matte smelting — Danger of a hot-top^ 
and its remedy — Amount of wind required — Comparative 
merit of Roots and Baker blowers — Relation between typo 
of furnace and volume of blast needed — Experience with 
forehearths at Leadville — Use of speiss in the furnace 
charge — Breaking lump speiss — Speiss roasting — Danger 
of poisoning workmen in speiss roasting — Size of water- 
jackets at the Arkansas Valley works 1 

Chapter II.— Extraction of Gold and Silver from Matte. 
The Hunt & Douglas process— Method employed at Belle- 
ville, 111. — Separation of gold from copper bottoms at 
Argo, Colo. — An experience at Leadville 15 

Chapter III. — The Calculation of Furnace Charges. 
Charge used at the Arkansas Vail oy works, Leadville — 
Example of a slag calculation — Relation between matte 
and slag assays — Charge used at Aguas Calientes, Mexico 
— Weighing up furnace charges — Best type of slags ... 18 

Chapter IV. — Types of Furnaces. 

Adaptability of certain slags to certain furnaces — Best 
position of matte-tap — Furnaces with outside forehearths 
— Furnaces with internal crucibles — Method of tapping 
matte — Type of furnace at San Luis Potosi, Mexico — 
Smelting converter slag and refuse 24 

Chapter V.— Spouts, Settlers, and Jackets. 

The Schumacher spout^ — The Hixon spout — Design of tap- 
jacket — Other kinds of slag-spouts; their defects — Kind 
of settler used at Anaconda, Mont.— Settlers at Aguas 
Calientes, Mexico — Difference in matte and slag separa- 

(iv) 



TABLE OF CONTENTS. 

tion in large and small settlers — Arrangement of water- 
jackets — Advantages of a small number of Jackets — Size 
and number of tuyeres — Proper height of jackets — Protect- 
ing a brick shaft with water-pipes — Tuyere-bags 28 

Chapter VI.— Blowing-in and Barring-down a Furnace. 
Blowing-in a lead furnace — Barring-down — Effect of wall 
accretions on the furnace-running — Running down charges 
preparatory to barring-down or blowing-out— Barring into 
the crucible — Blowing-in a copper-matte furnace — Im- 
portance of a heavy matte-fall in blowing-in 37 

Chapter VII. — Handling Blast-Furnace Slag. 

Method employed at the Arkansas Valley works, Lead- 
ville — At the Omaha & Grant works, Denver — At the San 
Luis Potosi works, Mexico — At the Pueblo Smelting & 
Refining Co.'s works, Pueblo, Colo.— At the East Helena 
works, Montana — Handling slag from copper-matte fur- 
naces—Amount of water required for granulating slag . . 42 

Chapter VIII.— Design of Lead Blast Furnaces. 

Tendency to increase height of furnaces — The Aguas 
Calientes plant — Strange result of the first smelting 
campaign at Aguas Calientes— Effect of bell and hopper 
feed on a lead-smelting furnace — Deductions from the 
Aguas Calientes experience — Comparison between stack 
and top-feed furnaces — Methods of feeding lead and 
copper blast furnaces 47 

Chapter IX.— Lead Slags and Losses in Lead Smelting. 
Example of a lead-furnace charge — Types of lead slags — 
Characteristics of high-lime slags — Losses of lead and 
silver in smelting — Losses in preliminary roasting-slag- 
ging — Percentage of fuel required in smelting — Effect of 
different kinds of coke 54 

Chapter X. — Improvements in Roasting Furnaces. 
Proposed combination of roasting, smelting, and convert- 
ing plants — The O'Harra furnace — The Hixon roasting 
furnace 59 

Chapter XI. — Smelting Raw Concentrates with Hot Blast 

AT Anaconda. 
Construction of stoves — Temperature of blast — Unsatis- 
factory results of the experiment at Anaconda 65 

Chapter XII.— Copper Converting at Anaconda. 

Difference in conditions governing copper converting and 
the Bessemer steel process — Reactions in copper convert- 
ing — Inapplicability of basic linings to copper converters 



VI LEAD AND COPPER SMELTING. 

— Result of an experiment on protecting the converter 
lining with water coils — Experiment with a basic lining — 
Amount of silica required for a converter charge — Com- 
position of converter slags — Old practice of relining con- 
verters at the Anaconda and Parrot works, Butte, Mon- 
tana — Composition of lining material — Cost of converting 
in the first Anaconda plant — Design of a converter plant 
— Cost of converting — Losses in converting at Anaconda, 66 

Chapter XIII.— Blowing a Converter Charge. 

Condition of charge indicated by flame coloration — Danger 
in over-blowing — Slag skimming — Practice at Aguas 
Calientes — Regulation of temperature in the converter — 
Management of the tuyeres — Indications of a finished 
blow — Method of developing heat in a cold charge — Dif- 
ference in amount of heat furnished by high-grade and 
low-grade matte — Use of silicious ore for converter lining 
at Aguas Calientes— Kinds of finish on converter copper 
— Distribution of silver in converter bars— Sampling cop- 
per bars for shipment 74 

Chapter XIV.— Design of Converter Plants. 

The Parrot works, Butte, Montana — Arrangement of the 
new Anaconda works — Dimensions of copper converters 
— Objection to very large converters — Mechanical devices 
for handling converters — Method employed at Aguas 
Calientes — Moulds for casting converted copi)er — Coj^e 
and core apparatus for making copper moulds 84 

Chapter XV. — Lining a Converter. 

Effect of lining material on composition of converter slag 
— Preparation of lining material at Anaconda — Method of 
lining at Anaconda — Life of a converter lining — Experi- 
ment on patching linings at Anaconda 94 

Chapter XVI.— Casting Anodes Direct from Converters. 
Trial at Anaconda — Effect on electrolyte — Method of 
casting — Difficulties to overcome — Grade of converter 
copper — Casting anodes from converters at Great Falls, 
Montana 97 

Chapter XVII.— Cost of Producing Copper at Anaconda. 
Details of estimate— Losses in various processes of treat- 
ment—Details of copper-converting plant at Anaconda, 99 

Appendix. 

Specifications of buildings and machinery for copper- 
converting p)lant. Anaconda INIining Co., Anaconda, Mon- 
tana, and accompanying working drawings 102 



LIST OF ILLUSTRATIONS. 



Fig. 1. Furnace and forehearth No. 1, Arkansas Valley 

smelting works, Leadville, Colo 3 

Fig. 2. Furnace and forehearth No. 1, Arkansas Valley 

smelting works, Leadville, Colo., front view ... 3 

Fig. 8. Forehearth, furnaces ISTos. 2 and 3, Arkansas Valley 

smelting works, Leadville, Colo 5 

Fig. 4. Forehearth, furnaces No. 2 and 3, Arkansas Valley 

smelting works, Leadville, Colo., front plate .... 5 

Fig. 5. The Hixon slag-spout, cross-section 29 

Fig. 6. The Hixon slag-spout, longitudinal section 30 

Fig. 7. Plan and section of settler used at Anaconda .... 32 
Fig. 8. Plan of furnace and settler at Aguas Calientes ... 33 

Fig. 9. Design of water-jackets 34 

Fig. 10. Matte-pots used at Arkansas Valley smelting works, 

Leadville, Colo 42 

Fig. 11. Matte settling pots at Omaha & Grant works, Denver, 

Colo. ; general view 42 

Fig. 12. Matte settling pots used at Omaha & Grant works, 

Denver, Colo. ; longitudinal section 43 

Fig. 13. Matte settling pots used at Omaha & Grant works, 

Denver, Colo. ; cross-section . 44 

Fig. 14. Slag-truck used at Omaha & Grant works, Denver, 

Colo 45 

Fig. 15. Slag-truck used at Omaha & Grant works, Denver, 

Colo 45 

Fig. 16. Proposed combination of roasting, smelting, and con- 
verting plant, elevation 60 

Fig. 17. Proposed combination of roasting, smelting, and con- 
verting plants, plan 61 

Fig. 18. The Hixon roasting furnace, cross-section 62 

Fig. 19. The Hixon roasting furnace, details of carriage and 

plows 63 

Fig. 20. Converters at Aguas Calientes 85 

Fig. 21. Converters at Aguas Calientes, details 87 

Fig. 22a. General arrangement of 8-foot copper converter ... 88 
Fig. 226. General arrangement of 8-foot copper converter ... 89 
Fig. 23. Cope and core apparatus for making copper moulds . 90 

(yii) 



Vlll 



LEAD AND COPPER SMELTING. 



Drawings of the copper-converting plant of the Anaconda Mining 
Co. accomi3anying appendix. 

Plate I. General arrangement, 

Plate IE. General plan. 

Plate III. Converters. 

Plate IV. Details of converters. 

Plate V. Details of trunnion ring, etc., for converters. 

Plate VI. Details of tuyeres and tuyere-boxes for converters. 

Plate VII. Details of converter-stand, etc. 

Plate VIII. Details of foundations for stands. 

Plate IX. Details of hydraulic cylinder, rack and gear for 
converters. 

Plate X. Hydraulic four-way valve. 

Plate XI. Details of runners from cupola to converters. 

Plate XII. Details of converter flues. 

Plate XIII. Cross-section of buildings. 

Plate XIV. Details of copper cupola. 

Plate XV. Details of charging-fioor. 

Plate XVI. Water-jacket furnace. 

Plate XVII. General plan of blast pipe. 

Plate XVIII. Details of matte-carts. 



CHAPTER I. 

Copper Matte Smelting. 

The matting of ores without the presence of copper to serve 
as a carrier to collect the silver has been attempted on many 
occasions, but has only resulted in failure, the losses being 
too large under ordinary conditions to admit of treatment in 
this way. A small percentage of copper can be used with 
success, but if the resulting matte runs below 5 per cent, 
copper it is very doubtful if the process can be made success- 
ful. In cases where a large tonnage of ore is to be treated 
with a small amount of copper, it would be advisable to 
crush and roast a part of the matte produced and smelt it 
along with the charge to supply the copper needed. But 
in cases where the ores are sulphides, without the presence of 
either copper or lead to act as a carrier, better results can be 
obtained by leaching in localities where the conditions will 
not admit of marketing the ore to custom smelters. The 
reasons are that the tonnage of matte produced is not easily 
transported, and when sold has to pay treatment charges as 
well as allow for losses in subsequent treatment. If it is 
possible by the concentration of the matte to save enough 
copper to overcome these objections, then it becomes a ques- 
tion of costs to determine which is the most economical, and 
each particular case is an independent proposition and should 
be treated accordingly. 

Emergencies arise, and it may be necessary to convert a 
lead-smelting plant into a mixed one of lead and copper, or 
vice versa. 

In 1890 at the works of the Arkansas Valley Smelting Co., 
in Leadville, copper ores of sufficient quantity to justify 
separate treatment were received, and accordingly two of the 

(1) 



Z LEAD AND COPPER SMELTING. 

old lead furnaces were changed to copper furnaces by the 
very simple process of filling up the lead crucible and placing 
an overflow pot under the slag-tap, which pot was later 
replaced by a forehearth of the type shown in Figs. 1 to 4. 

The running of these furnaces was at the time a novel 
feature in Leadville practice, as prior to that time all the 
blast-furnace work had been lead smelting and no strictly 
copper smelting had been done. The copper ores were mainly 
Irom the Maid of Erin, with a considerable quantity of iron 
sulphides from the Mahala and Wolftone mines. 

The greater portion of the ores that could be used on the 
charge were sulphides of such a character as to require crush- 
ing and roasting, and, as a consequence, the charge was about 
70 per cent, calcines, including roasted matte from the lead 
furnaces known as furnace or lead matte, which, however, 
contained about 7 per cent, copper, the result of concentration 
of small amounts of copper from the ores fed to the lead 
furnaces. This material was consequently fine enough to 
make an excessive quantity of flue dust and have a serious 
effect on the running of the furnace. 

It might appear to the inexperienced that the size of 
material to be smelted would have no great effect on the cost 
of treatment, losses and tonnage of a furnace, but if any such 
persons should have to contend with the complications arising 
from an exceedingly fine charge, they would soon come to the 
belief that there is a very close relation between the size of 
the charge material within certain limits and the tonnage 
that can be put through. 

There was a twofold object in running the copper furnace ; 
first, to treat separately the lead ores and such ores as con- 
tained no lead and some copper, and, second, to desilverize 
an accumulation of foul slag of long standing which previous 
administrations had left on hand. This slag was in part the 
bottoms of pots or that portion of the slag next to the layer 
of matte, and of old make, and the other part was the shells 
resulting from the Devereaux pots in use on the lead furnace, 
of which six were in blast most of the time. 

It is due to state by way of explanation that lead was 
rather scarce in all the Colorado smelters from 1889 to 1891, 
and, as a consequence, the lead furnaces were running on a 



COPPER MATTE SMELTING. 




Scale M incli = l foot. 

Fig. 1.— Furnace and Foeehearth No. 1, 
Arkansas Valley SMEiiTiNG Works, LEADViiiLE, Colo. 








Scale 1 inch = 1 foot. 

Fig. 2.— Furnace and 1^ orehearth No. 1, 
Arkansas Valley Smelting "Works, Leadville, Colo.; front view. 



4 LEAD AND COPPER SMELTING. 

short allowance of lead, 8 per cent, to 10" per cent, on the 
charge, with a long allowance of zinc and magnesia, and the 
conditions were such that the slag assays were seldom below 
1.5 ounce Ag and more frequently 2 and 3 ounces Ag per 
ton. As a consequence of the high character of bullion — 
300 to 400 ounces — and low percentage of lead this was to be 
expected, and to remedy it as far as possible and to prevent 
an excessive silver loss the greater portion of the shells was 
fed into the copper furnaces. If it had not been for this 
addition of coarse, fusible material to the charge, the furnaces 
would have been even more difficult to manage than they 
were. Being only 36 by 84 inches, the furnace was not large 
enough to put through a tonnage that Avould keep a constant 
flow open, and, as a consequence, intermittent tapping, the 
same as on lead furnaces, was the practice. The slag com- 
position attempted was 34 Si02, 33 FeO, 20 CaO, but owing 
to the feeding of so much slag of unknown and varying 
composition the analysis of the resulting slag shoAved up 
somewhat incorrectly owing to the presence of some lead, and 
did not permit of the formation of as clean slags as can be 
made when the charge contains no lead. It is a well-known 
fact that in matte smelting lead is scorified, and, owing to its 
strong affinity for silver, will carry it into the slag, whereas if 
only copper and silver are present the slags will be much 
cleaner. 

The furnaces were arranged with a double tap in front, and 
every time the overflow pot was changed the lower tap was 
opened and the accumulation of matte in the furnace drawn 
off. The amount thus obtained would depend very much on 
the condition of the furnace, and ran from one to six or eight 
pots of clean matte before the slag would make its appearance. 
The condition of the furnace was mainly dependent on the 
amount of matte produced by the particular charge the 
furnace was running on. 

In cases of reconcentration before shipment, when the 
matte production was very heavy, perhaps 40 per cent, of the 
charge, the furnace would cut ont all accumulations of crust 
in the shaft and below the slag-tap and get into excellent 
condition. On the same composition of slag and percentage 
of fuel, with the regular charge on, where the matte produc- 



COPPER MATTE S:MELT1NG. 




Fig. 3.— Forehearth, Furnaces Nos. 2 and 3, 
Arkansas Valley Smelting Works, Leadville, Colo. 






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For K" Bolts 



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Fig. 4.— Forehearth, Furnaces Nos. 2 and 3, 
Arkansas Valley Smelting Works, Leadville, Colo. 



6 LEAD AND COPPER SMeItING. 

tion was necessarily small to effect a concentration of ten into 
one or more, the zone of fusion would travel up and the 
bottom of the furnace become crusted to such an extent that 
the crucible would hold not more than one pot of matte 
between the slag and matte-taps. 

As a consequence of small matte production and high 
concentration, it frequently happened that the matte-tap 
could not be opened until a charge was put on the furnace 
that would produce more matte, either by making lower grade 
or by feeding back some of the matte already produced. 

As a general thing less than 15 per cent, matte production 
in a strictly matting furnace will result in the tuyeres becom- 
ing hard and black and the rising of the zone of fusion to the 
top of the charge, which, of course, results in the fuel being 
consumed before it should, and the consequent loss of the 
heat escaping up the stack, as well as the inevitable loss of 
silver that will follow smelting in a furnace hot on top. 
Many furnaces have been, and doubtless are now, run on less 
matte production, but to run steadily at its maximum capac- 
ity is what is expected of a well-behaved furnace, and in order 
to do this it is essential to have the matte production in 
excess of rather than below 15 per cent. Likewise the 
production of matte in a lead furnace has a very marked 
influence on the minimum amount of lead that can be used 
on the charge. For example, it might be quite possible to 
run successfully on as low as 7 or 8 per cent, of lead, reinforced 
by a heavy production of matte, and altogether unxDrofitable 
to attempt it without the matte. Unless the corrosive effect 
of something beside slag is acting in the furnace the cold 
blast will chill the slag, resulting in crusts and the elevating 
of the zone of fusion. With the zone of fusion high above 
the tuyeres in a lead furnace the losses by volatilization 
become higher until the maximum is reached, when the entire 
amount is lost. 

To return to the subject in hand, the handling of small 
furnaces with intermittent slag-tap and with matte produc- 
tion as low as 10 per cent, or lower, it must be said that it 
requires a very careful handling on the part of the furnace- 
men, even allowing that the slags are of the best possible 
composition. To keep the zone of fusion down at the tuyeres 



COPPER MATTE SMELTING. 7 

is the essential condition of success, and to attain this is not 
entirely in the hands of the man who figures the charges. 
The charge as it comes to the furnace may be either too fine 
or too coarse to obtain the best results, and while in the first 
case nothing but the coarsest of coke should be used, in the 
latter the use of a sledge hammer on the large pieces will 
materially improve matters. 

The proper placement of the charge in the furnace is a 
matter of the greatest importance, and it is hard to decide 
where the more experience is required, on the feed floor or in 
handling the furnace below. 

In a case where the furnace temporarily becomes cold, as is 
likely to happen through the variation in matte production, 
it may become advisable to increase the fuel or put back a 
charge of straight matte in order to repair the threatened bad 
effect, and if the feeder does not know his business, or if the 
furnaceman does not properly attend to his duties, the result 
may be slag in the tuyeres, a disagreeable experience with 
sledge hammers for a time, and what is known as muscular 
metallurgy. 

When a furnace has from any cause, be it fine ore, a bad 
slag, small matte production, or any other fault, become 
dark at the tuyeres and hot on top, to know the quickest way 
in which to remedy the difficulty is what is required of the 
man in charge. If it is a copper furnace the easiest way is 
by putting on matte charges until the tuyeres get bright and 
the crusts or accretions have disappeared sufficiently so as 
to warrant the matte charges being taken off or reduced in 
number by putting on one matte charge to one, two. or three, 
or more, of ore charges. It frequently happens at shift 
change, especially early in the morning when the night men 
are tired and are doing their work in a somewhat lazy and care- 
less fashion, that the furnace is allowed to run down and get 
very hot on top. At such time the operation of the furnace 
assumes the character of pyritic smelting, the oxidation of 
sulphur being much greater than when the furnace is fed at 
its proper level, with the result that the matte production in 
proportion to the charge falls off* rapidly and consequently 
increases in grade of copper contents. The excessive loss of 



8 LEAD AND COPPER SMELTING. 

heat up the stack results in the furnace running cold, and 
when suddenly filled up again the consequence is bad running 
and probable freeze-up. 

It is not the purpose of the writer to refer only to ideal con- 
ditions where everybody connected with the furnace knows 
his business and attends to it properly, but to take from the 
experience of the past that which may be valuable to others 
in the future. Too much blast will result in driving the fire 
to the top of the charge, and, providing the furnace will 
stand it, a reduction of the blast will assist the matte charges 
in bringing the zone of fusion down. On the other hand, 
tonnage must be maintained, and in order to make a furnace 
run on a good charge above 100 tons per day it is necessary 
to drive the wind into it at a lively rate, and a No. 7 Roots 
blower will have to make 140 revolutions per minute to 
furnish blast enough. This amount of blast would be all right 
for a copper furnace 42 by 120 inches at the tuyeres, but 
would be too much for a smaller furnace on the same charge, 
and entirely too much for a lead furnace of the same size. 

It is natural to discuss at this point what is the approximate 
amount of wind to blow in in order to obtain the best results 
in copper and lead smelting. Taking the listed displacement 
of a Roots No. 7 blower at 65 cubic feet per revolution, and 
allowing 140 revolutions per minute for a copper furnace 42 
by 120 feet at the tuyere, we get 9,100 cubic feet per minute 
for a furnace with cross section of 35 square feet, or 26 cubic 
feet of blast to each square foot of furnace section. Under 
ordinary conditions 70 revolutions per minute would be about 
the proper amount for a lead furnace of the same size to run 
on and would give 13 cubic feet per square foot of furnace 
section. 

These figures are of necessity only for normal conditions of 
charge and tonnage, and if conditions arise for greater or less 
tonnage or to reduce the loss by volatilization, the amount of 
blast supplied has to be regulated to suit those conditions. 

Also, the blowers may through wear become leaky to such 
an extent that they will not deliver the calculated amount of 
air to the revolution, and, as a consequence, would have to be 
run faster. I have never personally tested my blowers to find 
out how much they would deliver per revolution and at 



COPPER MATTE SMELTING. 9 

varying pressure, except at Anaconda, where a No. 4 Roots 
was run with a closed outlet at 20 revolutions per minute and 
gave a pressure of four pounds to the square inch in the blast 
pipe, which would mean that if run against a pressure of four 
pounds it would deliver no air at all, or, in other words, the 
leakage at four pounds was 100 per cent. 

With Baker blowers the leakage is much higher, and I do 
not believe that they could be made to generate a pressure of 
more than two pounds before the leakage would be 100 per 
cent. This would indicate what experience is slowly teaching 
all blast furnacemen — that the efficiency of Baker blowers, 
or blowers of that type, is much less at any pressure than 
blowers of the Roots type. Fifteen years ago, when Leadville 
was a new camp and when the writer received his first 
baptism of fire in the smelting business, there was not a 
Roots blower in any of the smelters that were in operation, 
and this tendency to oppose change has been maintained to a 
great degree even in the construction of new plants. It is an 
easily demonstrated fact that if the efficiency is lower the 
power required to deliver the blast to the furnaces must 
necessarily be larger than would be the case if blowers of the 
Roots type were used. As there are now two competing firms 
building this class of positive blowers, the writer can not be 
accused of unjust prejudice. 

The number of moving parts and the closer contact re- 
quired for a more positive delivery of the air, are the rea- 
sons for the greater efficiency and consequent lower consump- 
tion of power. The diflPerence in power consumption will 
amount to fully 50 per cent, in favor of the Roots type, and 
regardless of first cost would soon pay for entirely new blower 
plants for many works now in operation. The listed speed 
for a No. 7| Baker is only half that of a No. 7 Roots of the 
same displacement per revolution, so that on a basis of cost 
alone one Roots type is worth two Bakers, and on a basis of 
operation the ratio is two to one in favor of the Roots type, 
so that finally it would appear that the relative values of the 
two types is about four to one in favor of the Roots. 

The type of furnace has also much to do with the amount 
of blast required for successful running. If, for instance, it 
is a constant discharge, the flow of slag must be rapid enough 



10 LEAD AND COPPER SMELTING. 

to keep the slag-spout open, and this can not be done on much 
less than 60 tons of furnace charge per day. If the furnace is 
small and the character of charge such as not to admit of rapid 
smelting, it may be better to run an intermittent tap with a 
moderate blast than to try forcing the tonnage with high blast 
and run the risk of driving the fire up and running the 
furnace cold at the tuyeres. 

These were the conditions at Leadville. The ore was fine, 
the matte production small, and the tonnage that could be 
handled not enough to keep a constant discharge open. A 
forehearth of the Herreshoff type with water- jacketed sides 
was constructed, but when the furnace was started it was 
found that the cooling eflPect of the water was too great and 
the entire contents of the hearth soon became solid. The 
hearth was changed and the water-jackets allowed to remain 
empty with better results, but the neck connecting the hearth 
with the furnace would become closed and slag would be 
forced out of the tuyeres after about twelve hours' run, and 
it was necessary to change the hearth at least that often. 

The hearth was then constructed of cast plates with a pipe 
cast around the hole which connected with the furnace, and a 
similar plate with water circulating around the hole in the 
furnace front. This construction worked satisfactorily as 
long as the matte production was small, but if as much as 20 
per cent, matte was produced it would cut out and discharge 
the contents of the furnace on the floor. This type of furnace 
was not adapted to steady running under varying conditions, 
since the changing of the hearth necessitated a shutdown 
which was unavoidable and frequently of an hour's duration. 

After running some time under the conditions stated, with 
a matte production of about 10 per cent, and varying in 
grade from 15 to 40 per cent. coi3per, and in silver from 150 
to 600 ounces, the experiment was tried of mixing in a small 
percentage of speiss with the ore before roasting. At first 
the quantity used was one-fifth of the calciner charges, and 
as no bad effect was observed on the calciners or blast furnace 
the proportion was increased until one-half of the calciner 
charges Avas arsenical speiss that had been produced several 
years before and was regarded as a waste product. 

This speiss contained an average of 18 ounces Ag, one-tenth 



COPPER MATTE SMELTING. 11 

ounce An, 2 to 5 per cent. Pb, 15 to 20 As, about the same 
amount of sulphur, and the remainder iron. It had been 
made when the ores of the camp were chiefly of carbonate 
or oxidized character, and was a great annoyance at the 
time of its production as well as entailing a heavy loss of 
silver and gold in smelting. The accumulation of ten years' 
operations was stacked up in one corner of the slag-dump in 
the hope that by some undiscovered process it might some 
day be treated and the silver and gold recovered. There were 
about 3000 tons of it at this time on the dumps of the Arkansas 
Valley works, representing a large sum of money in gold and 
silver as well as the value of the iron excess as a flux for silica. 
At first this experiment did not promise to turn out well, as 
only the finest portions of the old pile could be used, and it was 
thought that Avhen they were consumed the golden eggs would 
be gone. It will be understood that the crushing of speiss in 
lumps, varying in weight from 20 to 80 pounds and of a hemi- 
spherical shape, is attended Avith great risk of breakage to 
the machinery, and especially when there is some metallic 
lead to be found associated with it. The speiss was therefore 
thoroughly culled before sending it to the crushing mill, and 
all large pieces broken in two or more pieces by sledging to 
avoid as far as possible the risk of getting lead into the 
crusher. 

But in spite of these x)recautions the machinery was fre- 
quently stalled, belts thrown ofl", and both belts and bolts 
broken, and finally the frame of a 9 x 15 Blake's crusher 
was split down through the jaws. The shaft was also badly 
bent and had to be straightened and rebabbitted in the 
bearing where the great strain had squeezed out the babbitt. 
But by this time rapid strides were being made in the con- 
sumption of the speiss pile that had been an eyesore for so 
many years, and after being encouraged by a 5 per cent, 
silver gain in one month's run over and above the silver 
charged in the smelting returns, it was seen that the heavy 
repairs to several crushers could be afforded. 

There was no attempt to keep the records of the copper 
furnaces separate from the lead smelting, since the copper 
smelting was carried on with a view of desilverizing dirty lead 
slags in combination with straight matting for direct profit. 



12 LEAD AND COPPER SMELTING. 

SO that the 5 per cent, gain represented a total of 15,000 
ounces on a charge of 300,000 ounces, and also no loss on the 
silver charged, which under favorable conditions would have 
amounted to about 9000 ounces, so that at a low estimate the 
gain was about 24,000 ounces. The contents of the speiss 
were not charged to the furnace, and so long as the smelting 
returns continued to show in that unusual and satisfactory 
fashion it was deemed advisable to go on utilizing it. 

The crushing continued day and night ; slowly, to be sure, 
but fast enough to use the stock on hand in about five months 
— all too soon for the writer, for by that time we had become 
speiss hungry and the dump was thoroughly explored by 
tunnels in the hope of finding more buried treasure. 

Next in importance to the crushing came the roasting of the 
crushed product. This was carried on in reverberatory cal- 
ciners of the standard type, with hearths 14 feet wide by 70 
feet long, the speiss being mixed with crushed sulphides of 
copper and iron in the proportion best suited to driving off the 
maximum amount of arsenic and sulphur without becoming 
too fusible. It was found that if the speiss constituted too 
large a portion of the roaster charge it was likely to become 
plastic as it came near the firebox, and in that condition 
would stick to the hearth, as well as to the paddles and rabbles, 
and would naturally refuse to part with its sulphur and arsenic. 

After experimenting with different roaster charges it was 
finally found best to use one-third speiss and two-thirds ore as 
giving the most satisfactory results, and this charge was con- 
tinued until the stock was used up. The ore that was roasted 
with the speiss contained about 27 per cent. Si02, 25 per cent. 
Fe, 30 per cent. S, 7 per cent. Cu, and was admirably suited 
to the purpose, as the silica rendered the roaster charges more 
infusible than they would have been otherwise, and the 
cxDpper furnished the carrier for extracting the silver in the 
blast furnace. The roasters were charged four times each 
shift with 3000 pounds of the mixture. 

The charge was worked the same as if it had been all ore, 
rabbled every twenty minutes and moved forward four times 
each shift of twelve hours. With five roasters running, each 
one handling twelve tons per day, we roasted twenty tons 
of speiss and forty tons of ore down to about 5 per cent, sul- 



COPPER MATTE SMELTING. 13 

phur. The arsenic was partially driven ojff, about one-tliird 
the original amount remaining in the roasted product, which 
I am inclined to believe was oxidized to a considerable extent. 
The charges were drawn from the furnaces into slag-pots and 
dumped on to a cooling floor. In order not to alarm the men 
unnecessarily and at the same time avoid danger of poisoning, 
the xDroportion of speiss to ore on the charge was increased 
gradually from 10 per cent, at first, to 15 i)er cent., 20 per 
cent., 30 per cent., and 40 per cent, of the charge. But with 
so much as 40 ]3er cent, or 50 per cent, the fumes, when 
drawing the charges, became dangerous, which, with the dis- 
advantage of the plastic condition referred to previously, 
made it both necessary and desirable to decrease the amount 
of speiss. Some of the men were badly poisoned about the 
nostrils, and, in fact, one was dangerously ill from the effects, 
but on the assurance of the company's physician that this 
could be guarded against, the work was kept up. A prepara- 
tion of hydrated oxide of iron was put up as a salve and fur- 
nished to the men to put into their nostrils, and whether this 
preventive was the cause or not, we did not have any further 
serious trouble or fatalities. 

The calciners and the calcined material, after cooling, were 
covered by a white sublimate of arsenious acid, which had the 
appearance of frost on a cold morning. After cooling from 
twelve to twentj^-four hours the roasted material was loaded 
on cars and sent to the blast furnace, where it was smelted 
with silicious copper ores from the Sedalia mine, Colorado, or 
from the Eureka Hill in Utah, together with the addition of 
enough raw sulphides from the Leadville mines to keep the 
matte production at such a ratio that the furnaces would 
run as regularly as possible. As before stated, these furnaces 
Avere originally lead furnaces, with the crucible filled with 
brickwork and silicious lining. 

The jackets were of the ordinary cast-iron type in use in all 
lead smelters and of the ordinary height, i, e., about four feet 
six inches. This height of jacket does very well for a lead fur- 
nace where the blast is not high enough to raise the zone of 
fusion, but in the smelting of these ores there was great trouble 
from burning out of the brickwork over the jackets. The fur- 
naces would treat about 40 tons of ore per day on a 16 per 



14 LEAD AND COPPER SMELTING. 

cent, fuel consumption, and, in addition, 20 tons of slag. It 
would naturally be expected that smelting with so much 
speiss on the charge a considerable quantity of speiss would be 
produced and would separate from the resulting matte. But 
such was not the case. When the furnaces were tapped it 
would frequently spark in the way which is characteristic of 
speiss, but after cooling there would be no line of separation 
in the pots, and upon being crushed and roasted and resmelted 
the product was a matte of very clean appearance with 40 to 50 
per cent, copper, the arsenic contents of which did not exceed 
5 per cent. 



CHAPTER II. 

Extraction of Gold and Silver from Matte. 

The matte produced at the Arkansas Yalley Smelting 
Company's works was shipped to the refinery at Argentine, 
Kansas, and treated by the improved Hunt & Douglas process, 
or reshipped to Block & Hartman, Belleville, 111., and there 
treated by a leaching process. 

In the Hunt & Douglas process the matte is roasted at 
very Ioav temperature, so that copper sulphate and oxide 
result without forming any silver sulphate. It is then 
leached with dilute sulphuric acid, the gold, silver, and lead 
remaining in the residue. The copper solution is chloridized 
by the addition of chloride of lime and the copper precipitated 
as subchloride by passing sulphurous acid through the solu- 
tion. Ilie subchloride of copper is reduced to suboxide by 
milk of lime, whereby chloride of calcium for further use is 
recovered, while the suboxide of copper has only to be reduced 
to ingot by a simple smelting. 

The Block & Hartman process for recovery of gold is some- 
what similar to the practice at Argo, Colo., if not identical 
with it. The matte is first roasted to convert the silver to 
sulphate, when it is leached out with water and precipi- 
tated on metallic copper, the gold remaining behind in the 
matte. Or it is concentrated to black copper in the rever- 
beratory furnaces, granulated, ground, roasted, and leached 
with salt solution by the Augustin process and the silver 
precipitated in the usual manner. 

The residue containing the copper and gold is then concen- 
trated in a reverberatory furnace until a small amount of 
copper is extracted as a copper bottom, carrying nearly all 
the gold. The subsequent treatment and separation of the 

(15) 



16 LEAD AND COPPER SMELTING. 

gold from these copper bottoms has been much talked of 
among metallurgists, and is supposed to be a secret of such 
importance that Mr. Pearce, of Argo, though generally most 
liberal in imparting knowledge, is for business reasons unable 
to divulge it. 

An experience that occurred while running the furnaces at 
Leadville bears on this point, and the reader may be left to 
form his own conclusions about the separation of gold from 
copper bottom^s at Argo. In concentrating the matte which, 
owing to the large percentage of lead slag on the charge, 
contained a small amount of lead, there were formed some 
bottoms which, while they could not be called properly copper 
bottoms, will answer for an illustration. These bottoms were 
metallic in appearance and would ring like bell metal when 
struck with a hammer, were quite malleable, and could not be 
broken. The matte, which separated from them very easily, 
did not contain more than 55 per cent. Cu, so that without the 
intermixture of lead they would not have been formed. 
Having them on hand it became a question what to do with 
them. An assay showed that the gold had almost all left the 
matte and gone into these bottoms. The experiment of fusing 
a portion in a scorifier with the addition of a pyritous ore 
containing no silver and a little gold was then tried. It was 
found that the gold continued to concentrate in the portion 
which remained metallic, while very little, if any, went into 
the matte formed by the pyrites on the surface of the fused 
charge. 

In fact, there was a scorifying action, with sulphur as the 
agent of concentration, and the resulting metallic portion all 
the time growing richer in gold as it got smaller. This action 
was kept up until the greater portion of the copper had been 
removed, and the resulting button could be easily cupelled with 
lead whereby the remaining copper was removed, and a gold 
button remained which could be dissolved and precipitated as 
fine gold. Certainly the gold can be extracted from copper 
bottoms in this way, and since it is a well-known fact that mill 
concentrates and even tailings are to be had from certain gold 
mines around Black Hawk, Colo., with which the excess copper 
could be reconverted into matte and the gold concentrated into 
such, a small amount of copper as to admit of its being refined 



EXTRACTION OF GOLD AND SILVER FROM MATTE. 17 

it remains for the reader to decide whether it is a secret 
process or only one of the tricks of the trade. It is at all 
events only applicable to just such conditions as exist in this 
process of extracting the silver first and subsequently the 
gold, and it is difficult for anyone to see how a competitor 
could possibly take advantage of the general knowledge of the 
process, especially as the electrolytic method is better and 
cheaper. 



CHAPTER III. 



The Calculation of Furnace Charges. 



The calculation of a slag for the furnace is illustrated by 
the following approximate make-up of the charges we were 
using in matte smelting at the Arkansas Valley works, as 
described in the previous chapter, taking the analysis of the 
ores from memory. They may not be exact, but they are 
close enough for the purpose of illustrating the method ordi- 
narily used by metallurgists for figuring blast-furnace charges. 





1 


o 


O 


d 

o 


i 


d 

OS 
O 


? 


m 




3 


3 




^ 


02 


4i 


Uh 


4i 


U 


■J 


CO 


^ 


o 




-t^ 


a 


02 




M 


a 


m 


a 


CO 


a 


M 




xi 


Ol 


-o 


a) 


-o 


a> 


'0 


(D 


-o 


© 


-d 




tD 


o 


C! 


o 


a 


o 


a 


O 


a 


o 


a 




<V 


u 


S 


u 


s 


u 


s 


u 


s 


tn 


a 




^ 


© 


o 


® 


o 


® 


o 


<x> 


o 


© 


o 




Pn 


cu 


Ph 


Oh 


Ph 


Oh 


Ph 


Ph 


Ph 


0. 


Calcined ore and speiss 


500 


20 


100 


42 


210 








5 


25 


5 


25 


Raw sulphide ore 


150 
150 


26 
60 


39 

80 


32 
14 


48 
21 


1 

6 


1.5 
9 


30 
2 


45 
8 


8 
3 


12 


Silicious ore 


4 


Lime rock 


200 


3 


6 






52 


104 




73 








1000 




225 




279 




114 


41 



I 



There is no algebraic mystery or X, 1^, and Z equation to be 
solved by higher mathematical methods than percentage, and 
the results are just as reliable. There is generally a great 
deal of mystery thrown about this portion of the metal- 
lurgist's work, as if there were a fear that if the younger 
members of the profession were put in possession of the com- 
bination they might soon be competitors. 

The writer has no hesitation in saying that the calculation 
of a charge, when it has been decided what ores are to be 
smelted, is the simplest thing about the work. The difficulty 
is more often to decide what to put on and how much matte 
to make to have it run to the best advantage. There are cer- 

(18) 



THE CALCULATION OF FURNACE CHARGES. 19 

tain things a metallurgist must know ; for instance, he must 
be able to tell approximately how much matte the charge is 
going to make and how much of the iron is going into it in 
order to make the proper deduction from the total amount on 
the charge. The writer has been accustomed to have all 
analyses of ores, whether sulphides or oxides, determined or 
figured as FeO, and then, knowing from experience and from 
the running of the furnace about the grade of matte that cer- 
tain charges will produce, he proceeds to multiply the total 
pounds of copper by a certain factor to get the weight of matte 
that the charge will produce. 

In the case of the charge above cited let us assume that the 
matte will assay 25 per cent. Cu. He accordingly multiplies 
the 41 pounds of copper in the charge by four, which gives the 
matte production at 164 pounds. A 25 per cent. Cu matte 
with the amount of impurities, such as Pb and As, that would 
naturally go into it from ores of ordinary character, would 
contain the equivalent of 40 per cent. FeO, and 40 by 164 
equals 65.9 pounds FeO, to be deducted from the total of 279, 
and leaving off the tenths gives 214 pounds FeO to go into slag. 
The sum total of the Si02, FeO, and CaO, after deducting the 
iron that will go into the matte, is 553 pounds, which according 
to slag determinations is generally 90 per cent, of the slag. 

Accordingly 553 pounds of Si02, FeO, and CaO, when put 
into slag together with AI2O3 and other oxides, would repre- 
sent approximately 620 pounds slag to a 1000-pound charge. 
These 620 pounds divided into the amounts of Si02, FeO, and 
CaO that would go into the slag would give a calculated 
analysis of 36.3 per cent. Si02, 34.5 per cent. FeO, and 18.4 
per cent. CaO. 

In explanation of the reason why it is assumed that the 
matte produced will assay 25 per cent. Cu, it may be stated 
that this would be the experience with ordinary running of 
the furnace on the charge figured, the excess of the sulphur 
being burned off. The grade of matte being dependent on 
the condition of the furnace, the depth of charge, and the 
amount of blast used, a knowledge of the effect of these con- 
ditions is necessary in order to make this assumption. 

As regards the percentage of Fe in different grades of 
matte, that is a matter of local experiment depending on the 



20 LEAD AND COPPER SMELTING. 

amount of other impurities that the matte can absorb from 
the ores on the charge. For example, if the charge contained 
no arsenic, antimony, or zinc, the iron contents of a 25 per 
cent, copper matte resulting from smelting them would be 
higher than if these impurities were present in the ores. If 
present they will go partly into the matte and will displace 
iron. 

Roughly speaking, matte is a compound of the sulphides, 
arsenides, and antimonides of the metals, while slag is a union 
of the oxides of the metals with the oxide of silicon. The 
characteristics of either matte or slag as regards melting 
point, conductivity, and specific gravity are capable of as 
many variations as the composition. As in the case with the 
alloys of the metals themselves, where by certain combina- 
tions it is possible to produce one that will melt at a 
much lower temperature than any of its constituents, so it 
is with slags and mattes, but in a less marked degree. There 
are certain combinations of Si02, FeO, and CaO that form the 
more fusible slags, and are adapted for certain purposes. 

The science of metallurgy is the application of this knowl- 
edge to the formation of fusible compounds in order to assist 
in the recovery of the metals or their sulphides, arsenides, or 
antimonides. In the case of the charge figured above the slag 
would be silicious enough to insure the driving off of a reason- 
able amount of sulphur and arsenic and still near enough to a 
neutral slag to flow freely, and in view of the grade of 
matte produced should be clean or at least as low as 1 per 
cent, of the matte assay. 

The assay of slags is a thing as much dependent upon the 
means provided for settling them and the care with which 
they are handled as upon their composition. Still it bears a 
very close relation to the assays of the matte or bullion, 
which are closely related to each other, subject to varying 
conditions. Generally speaking, the slag will assay 1 per 
cent, of the matte in copper work and 2 per cent, in lead 
work. In the latter case the matte will, as a rule, contain 
about one-fifth as much silver per ton as the bullion if the 
amount of matte is normal ; but if the matte production is 
large compared to lead, the relation between assays will fall 
lower. In a case where 400-ounce bullion was being pro- 



THE CALCULATION OF FURNACE CHARGES. 



21 



duced on a 11 per cent, lead charge with 5 per cent, of matte, 
the matte would assay approximately 80 ounces, and the slags 
if well settled, 1.6 to 2 ounces. But these things are so vari- 
able that hardly any rules can be given that will not have as 
many exceptions as applications. 

The calculation of a charge fcr the copper furnaces at Aguas 
Calientes, Mexico, is here given : 







O 


O 


d 


i 




i 


02 






fl 






+i 


OQ 


<J 


fe 


^ 


o 


«• 


OQ 


■J 


o 




-tJ 


S3 


m 


c5 


m 


a 


09 


a 


00 


fl 


aa 




^ 




T3 




•CJ 


0) 


'O 




"5 




•o 




bc 


o 


d 


o 


d 


o 


d 


o 


d 


o 


d 




© 


^ 


s 


u 


3 


t-i 


s 


h 


fl 




d 




^ 


a> 


o 


01 


O 


<s 


p 


© 


o 


© 


o 




Ph 


^ 


fu 


Ph 


Ph 


Ph 


Pm 


fe 


Ph 


PM 


General mixture 


500 


25 


100 


30 


150 


7 


35 


20 


100 


6 


30 


Mixture 


200 


20 


40 


36 


72 


2 


4 


33 


66 








Silicious ore 


200 
100 


50 
20 


100 
20 


7 
19 


14 
19 


3 
4 


6 
4 


4 
22 


8 
22 



19 





Copper ore 


19 


Converter slag 


150 


25 


37 


62 


93 














5 


7 




260 
90 


3 
19 


8 
17 






52 
2 


135 

2 


"21 


* 18 


id 




Concentrates 


32 


29 


9 




1500 




322 




877 




186 ). . 


214 




65 



Assuming a 25 per cent. Cu matte, 65 pounds Cu X 4 = 260 
pounds matte X 40 per cent. = 104 pounds FeO to be deducted 
for matte, leaving 273 FeO to go into the slag. 322 Si02, + 
273 FeO, + 186 CaO = 781 -^ 9 = 870 pounds slag to the 
charge. Dividing 870 into 322 Si02=37 per cent.; into 273 
FeO = 31.3 per cent. ; into 183 CaO = 21.4 per cent. 

In the figuring of charges the estimation of the Si02, FeO, 
and CaO contents to the one-tenth of pounds is a fineness of 
calculation that is totally wasted on the man with the shovel 
at the charge scale. Besides, many other conditions prevail 
which do not warrant the assumption that the ores are uni- 
formly of the same composition. The bedding may not be 
regular, and when a new bed is opened more of the top layers 
go into the charges than later, and if the bed is in a square 
bin, as the opposite side is approached, the top layers are 
exhausted and the bottom layers alone will for a time take the 
place of the entire bed. In the face of such conditions cal- 
culations that may have been perfect melt into nothingness, 
and it becomes a necessity to be able to tell the character of 
the slag as it comes from the furnace and make such correc- 
tions as are necessary, frequently without analysis. 



22 COPPER AND LEAD SMELTING. 

The larger the ore mixtures the more evenly the furnaces 
will run, but if the mixtures are small and it is necessary to 
change the charge every day, there is not sufficient time to 
settle down to business and get the charge corrected to just 
the proper point to do the best work. The weighing of the 
charges is a part of the operation that requires the constant 
attention of a careful foreman. Owing to the scattering of 
fine ore over the scales while shoveling in from the bins to the 
car or barrow, the scales should be swept clean after each 
charge and the accuracy of the balance frequently verified. 

The character of slag to be run on is largely a matter of 
personal preference with different metallurgists, but in 
general it is determined by the local conditions regulating the 
cost of iron and lime. 

The limits within which the experience of the writer has 
reached the best results are for copper blast-furnace work : 

SiOs, 30 to 38 per cent. ; FeO, 30 to 40 per cent. ; 
CaO, 10 to 25 per cent. 

But under ordinary conditions the slag that is best adapted 
to a constant discharge and regular run is : 

36 per cent. Si02, 33 per cent. FeO, 21 per cent. OaO. 

It may change considerably either way and continue to run 
just as well, and a slight correction will bring it back to the 
original type. 

I do not pretend to say that slags can not be made higher in 
Si02 than 38 per cent., but they run very slowly compared to 
slags with less, and the coke seems to burn out faster than 
the charge smelts, leaving the furnace full of cold stuff, and 
accretions form rapidly at the tuyeres where the blast strikes 
the slag. 

It has been stated in other works that the composition of 
the slag may vary without serious results within much wider 
limits than those stated by the writer. While this is true of 
furnaces where the slag is tapped periodically and can collect 
in the furnace and run out with a rush, it is not true of a 
constant-flow furnace, as any such slags would solidify in the 
spout, and before the furnace could be opened the accumula- 
tion in the furnace would run out from the tuveres. At 



THE CALCULATION OF FURNACE CHARGES. 23 

Anaconda, where the charge was largely slag from the copper 
converters, I have often had the analysis show 40 8102, 43 
FeO, 9 CaO when the matte production was large (about 30 
percent, of charge), but in spite of this the furnace would 
run much slower and the spout would be difficult to manage. 
It was found that when the Si02 was near 36 per cent, and 
the sum of the FeO and CaO about 54 per cent., the best 
results w^ere obtained. This was also the experience with 
the copper furnaces at Aguas Calientes, and while it might 
appear to be a better commercial proposition to make a more 
silicious slag, still, if the furnaces will not smelt enough ore 
on that kind of slag to supply the material to keep up the 
circulation in the spout and prevent its freezing, it does not 
pay to run on such slags. 



CHAPTER IV. 

Types of Furnaces. 

The type of furnace has much to do with the kind of slags 
that can be run in it. An intermittent tap will handle slags 
that would not do at all for a constant flow, for the reason 
that in the latter case the slag would chill so quickly that the 
taphole would be choked up, causing shutdowns that in many 
cases would develop into freeze-ups. The best type of inter- 
mittent tap is a single slag-tap in front and matte-tap on the 
side at about one foot lower level. The hearth should incline 
forward from a foot below the level of the tuyeres at the 
back to two feet in front. This is a better arrangement than 
having the matte-tap immediately below the slag-tap in front, 
for the reason that the matte can be tapped at any time while 
the slag is running without removing the settler. The con- 
stant-flow furnace, on the other hand, gives better results 
when the matte and slag are run into a forehearth of from 5 
to 10 feet in diameter and are given ample time to separate. 
The volume held in ordinary overflow pots is too small for 
complete separation. Like the deposition of sediment from 
water, matte and slag must come to a standstill in order 
to separate thoroughly. 

A forehearth requires considerably more tonnage from the 
furnace to keep it open than the shallow crucible and over- 
flow pot, and also as the settler is made larger more matte on 
the charge ; otherwise the radiation is enough to form a very 
thick crust of slag on the top and sides and greatly diminish 
the size of the settler. The interior of the settler takes the 
shape of a pear with the larger end down, where the corrosive 
action of the matte is the greatest, w hile in the upper portion 
the slag once chilled remains there. 

(24) 



TYPES OF FURNACES. 25 

The handling of furnaces with internal crucible is attended 
with about the same amount of difficulty as those with fore- 
hearths, the difference being that in the first case, if it is im- 
possible to tap the matte from the furnace, it can be run out 
into the settler in front until such times as the accumulation 
in the furnace renders the breast soft enough to drive a bar. 
But if the matte- tap on the forehearth is lost through failure 
to put in a bar immediately after tapping matte and while it 
is yet soft, or through chilling of the contents owing to slow 
running of the furnace, then all is lost ; for while the furnace 
may be put into good condition in a few hours by matte 
charges, the settler being lost and too large to move, there is 
nothing to do but blow out the furnace and dig out the settler. 
This is the only bad feature about large settlers; every 
other point is in their favor, especially in connection with 
converters, as they will permit a sufficient amount of matte to 
accumulate for the converter charge. 

The separation between matte and slag is much better in 
large than in small settlers, but the weight of a settler 10 feet 
in diameter by 4^ feet deep, filled with slag and matte, is too 
great to move it unless extraordinary appliances are provided. 
In all cases after tapping matte, unless the breast is exceed- 
ingly soft, as soon as the flow has been stopped with clay, a steel 
bar should be driven in slowly until the point has just pene- 
trated the clay and entered the crucible or hearth. This is to 
provide against a hard breast when the next tap is to be 
made, and is a very important part to attend to in the run- 
ning of the furnace. If, for instance, it is neglected, as it is 
likely to be owing to the attention of the men being drawn 
to other things, it may mean several hours of sledging to 
again open it although the breast be reasonably thin. 

With the matte-tap immediately below the slag-spout, as 
was the case in Leadville, the slag was drawn off until the blast 
blew out at the taphole. The settler was then removed as 
quickly as possible and the bar driven into the matte-tap, a 
slag-pot run under the matte-spout and the bar withdrawn by 
means of the ring and wedges, which are absolutely necessary. 
All the matte in the furnace would be tapped out without 
stopping the flow, except for a few seconds to change pots. 
The stoppage was done by means of a pole consisting of a 



26 LEAD AND COPPER SMELTING. 

piece of iron shaped like a tinner's soldering-iron, about 15 
inches long and 2| inches in diameter, welded to a gas-pipe 
handle about 10 feet long. The heavy point would be inserted 
in the taphole and the stream interrupted until the full pot 
could be drawn away and an empty one substituted. 

The taphole in front is not a good arrangement. It should 
be on the side of the furnace but as near the front as possible. 
Its position in front at Leadville was unavoidable on account 
of the way in which the furnaces were crowded together. It 
frequently happened that after the settler had been removed 
and before the matte-tap could be opened, the furnace would 
be full of slag and the slag-tap have to be opened again to 
prevent it running out at the tuyeres, which it sometimes did. 
With the matte-tap on the side the work of tapping matte can 
go on at any time without interfering with the slag-tap or 
necessitating the removal of the settler. This point is worth 
mentioning, because if it is known how not to do a thing it 
may be of great assistance in finding out how to do it. 

At San Luis Fotosi, Mexico, they have a number of lead 
furnaces converted into matting furnaces by filling the cru- 
cible and putting in a sloping hearth of firebrick. The crucible 
is about two feet deep below the tuyeres in front, and a tap- 
jacket is put in on the side near the corner of the furnace, and 
high enough above the floor to admit a slag-pot under it, with 
slightly depressed floor space immediately around the tap. 
These furnaces are used for the concentration of matte made 
in the lead furnaces, whatever metallic lead there is produced 
coming out Avitli the matte and being recovered on the dumj). 

Running a blast furnace in connection with a converter 
plant of the size of the one at Anaconda, where the sole pur- 
pose is to work up refuse material and smelt over converter 
slag, is a much simi3ler operation than smelting ores which 
are widely different in character and only mixed mechanically 
in the charge. The fact that a portion of the charge has been 
fused previously, no matter what the composition of the fused 
material may be, has a very beneficial effect on the smelting. 
This is probably due to the fact that the fused material melts 
higher up in the shaft than the unfused and has a dissolving 
and uniting effect on the constituents of the latter. 

With a charge that was 60 per cent, converter slag, it was 



TYPES OF FURNACES. 27 

only necessary to add raw ore to make the matte sufficiently 
low to extract the copper from the slag and waste material, 
together with lime rock to assist in the fluxing of the surplus 
Si02. The FeO was already in the converter slag and in 
combination with Si02, so that such smelting can not be said 
to require any special mention from a scientific point of view, 
but merely as a matter of practice in connection with convert- 
ing to recover the copper in the slag 



CHAPTER Y. 

Spouts, Settlers, and Jackets. 

The form of spout and tap-jacket for a constant-flow furnace 
is a very important matter and one of which no special 
mention has been made in any work on the subject. The 
spout ordinarily in use in Montana is made of one-inch pipe 
put together as closely as possible, and in such a way as to 
form a channel of four and one-half feet long, semicylindrical 
in shape, with one end open and one partially closed. A 
piece of sheet iron is fastened to the curved outside, and clay 
or a mixture of ground quartz and clay is rammed in between 
the pipe and against the sheet iron. When in position with 
the open end against the front of the furnace (see drawing of 
Anaconda furnace) the spout allows the slag to flow in a 
steady stream into the forehearth without the escape of blast, 
the slag stream flowing over the partially closed end and 
thereby trapping the blast. This spout is known as the 
Schumacher spout, and has been in use for many years. It 
does good w^ork, but the weak point about it is that the quartz 
or clay is eaten out by the slag and matte, and frequently 
the contents of the furnace flow through some defective point 
in the spout. 

To overcome these defects and prevent runaways the writer 
had the coil cast into an iron spout of the shape shown in 
Figs. 5 and 6. It was found that fewer pipes were necessary 
for this kind of a spout than with the old construction, and 
the number of coils was reduced to four. 

The cast iron proved to be sufficiently cooled by the water 
circulating through the coils to keep the spout from being 
attacked by the matte, except at the tip, where the stream 
falls into the settler. At this point the cast-iron covering is 
eaten away and the pipe is soon laid bare. 

(28) 



SPOUTS, SETTLERS, AND JACKETS. 



29 



If the pipe used in making the spout is not of the best 
quality it may soon be eaten away at this point and the spout 
rendered worthless. This is the most satisfactory spout that 
the writer has ever used on a constant-discharge furnace, as 
the cooling effect of the water is reduced to a minimum and 
is only what is required to keep the spout from being eaten 
out by the matte. 



\ 

> 
1 




/^ 




o 

<M 


^^ 




n 


\ 


vV J 


') 


/ 


\ V ( ) o^ 


/ 


n ^ 


^^ ^ 


y 


u 








1 



Fig. 5.— The Hixox slag-spotjt ; cross-section. 

The pipe used may be of any size, say from j to 1| inches, 
according to the water pressure, but must be of the best quality 
and of extra thickness. The coil should be made very care- 
fully and put together with malleable ells and return bands. 
It should then be heated to drive off all grease or oils, and 
while warm painted with graphite mixed in benzine. Several 
coats of this should be put on and allowed to dry before the 
coil is put into the sand mould to receive the cast-iron cover- 
ing. The cast should be poured as cold as possible in order to 
avoid burning the pipe. 



30 



LEAD AND COPPER SMELTING. 



Any good foundryman should be able to make these spouts 
without much difficulty, but still it sometimes happens that 
the pipe will be plugged by being fused and great care is 
necessary. 



JJ 



«M 



11 



s 



>* N.NL 
X) INLET 



eoTTON plAte of Hubnac.e 



SPOUT & JACKET 
COUPLING 



w- 



SETTLER 
WALL 




Fig. 6.— The Hixon slag-spout; longitudinal section. 

It is best to have the tap-jacket made of bronze, but I see 
no reason why a good quality of cast iron, or especially malle- 
able iron, would not do, although it would be more liable to 
crack. There is a yoke cast on the front of the jacket immedi- 
ately surrounding the taphole which is cored out for water 
space. The inside dimensions of this yoke are the same as 
the outside dimensions of the spout, and the yoke is to prevent 
the escape of slag and matte from between the spout and 
tap-jacket. The yoke should be cast larger near the jacket 
and the spout dovetailed, so that when the two are put 
together they will fit tightly and need no bolts or other clamps 
to keep them from springing apart. 

Spouts have been made in other ways, some of boiler iron, 
flanged and riveted together in such a way as to form one of 
much the same shape as that described, but the trouble is 
that there is too much cooling effect on the stream of slag 
flowing from the furnace, the result being that a skull forms 
in the spout which grows thicker rapidly, and finally the 



SPOUTS, SETTLERS, AND JACKETS. 31 

stream is interrupted entirel}^ by the closing up of the channel. 
The furnace goes on smelting and the slag soon runs out of 
the tuyeres because it has no other outlet from the furnace. 
A spout with the coils, such as that described, exerts the least 
possible cooling effect on the slag stream, and if the furnace 
runs at any reasonable tonnage and the slag is within the limits 
of composition mentioned on a previous page, will keep open 
and give no trouble at all from freezing up. A fire, either of 
wood or lump coal, should be kept on top of the slag stream, 
and by this means the stream should be liquid and show no 
shell or crust from the furnace to the end of the spout. Spouts 
constructed of boiler iron are a constant source of annoyance 
and will freeze up when the furnace is running at its best. 

At Anaconda the settler was much smaller than at Aguas 
Calientes, and of a different tj'pe. (See Fig. 7.) The slag 
and matte were discharged together from the spout into 
the settler, where the matte by its greater gravity would go 
to the bottom and pass under the partition of pipe coil and 
discharge at a lower level than the slag at the opposite end. 
This partition and the lining of the slag end of the settler 
were originally made of brick, but it was found that when 
the furnace ran a heavy tonnage the partition would be 
destroyed as well as the lining in the slag end of the settler. 
Pipes with a water circulation were substituted and worked 
very well as long as the tonnage was kept up, but would chill 
the contents of the settler if it became low, or was shut down 
for more than twenty minutes without tapping out. It is in- 
teresting to note that the copper contents of the slag dis- 
charged from these small settlers at Anaconda were much 
higher than slags made at Aguas Calientes, where the settlers 
(Fig. 8) were the largest in use (10 feet in diameter). On a 
charge producing 55 per cent. Cu matte at Anaconda the Cu 
contents of slag would average nine-tenths per cent., while 
under the same conditions at Aguas Calientes the Cu would 
not be more than five- tenths per cent., showing that the 
capacity of the settler has much to do with the cleanness of 
the slag. With lower grades of matte the slag assays would 
be lower, but about the same relation existed in the two cases. 
The jackets for a copper furnace of large size can be arranged 
in many ways, though there are but few good designs. In 



32 



LEAD AND COPPER SMELTING. 



some cases the furnace is all one jacket from the crucible 
or base-plate to the feed-door, and the other extreme was 
embodied in the furnaces at Aguas Calientes, where there 
were twenty-four jackets on the furnace, counting the tap- 
jacket and spout. It is a very poor plan to multiply troubles 
unnecessarily in any business, and especially water-jackets. 




Fig. 7.— Plan and section of settler used at Anaconda. 

The greater the number the more opportunities there are for 
one to leak or the water connection to become clogged and 
allow the jacket to burn out. On the other hand, it is not 
advisable to make the jackets too large. The middle course 
is the best. The front and back should be of two jackets each 
(see Fig. 9). The sides may be divided into any number to 
suit the conditions, but so that there shall be no seams or 
rivets exposed on the inside of the furnace. Jackets become 
very expensive as the size increases, on account of the unusual 
width of sheets necessary for the inside. Better results can 
be obtained by dividing the side of a 120-incli furnace into 
three or four jackets of equal dimensions. With three jackets 
to the side and two tuyeres to the jacket, there are six 
tuyeres to the side and twelve tuyeres to the furnace, which 



SPOUTS, SETTLERS, AND JACKETS. 



33 



is quite sufficient, providing they are made 4 to 4^ inches in 
diameter. Four jackets to the side and two tuyeres to the 
jacket gives 16 tuyeres to the furnace, which is rather too 
many, but if they are made 3 to S^ inches in diameter they 




Fig. 8.— PliAN OF FURNACE AND SETTLER AT AGTTAS OAIiIENTES. 

will be quite satisfactory. The objection to a large number 
of tuyeres is that they will be so close together that the 
noses that form in front have a tendency to unite and form a 
dark band all the way around the furnace, causing the zone of 
fusion to travel up. It is a mistake to have the tuyeres point 



34 



LEAD AND COPPER SMELTING. 



downward, as is seen in some furnace drawings. The reason 
is that in case slag fills the tuyeres, as it is sure to do some- 
time in a long run, the slag can not escape and solidifies in 
the tuyeres, whereas with a simple conical-shaped tuyere put 
in with its axis perpendicular to the face of the jacket, the 
slag can escape until the tuyere can be plugged and the 
trouble rectified. 




Fig. 9.— Design of water-jackets. 

It is also a mistake to put the jackets in a double row, 
one above the other, for the reason that whatever mud or 
scale there is in the water will settle in the bottoms of the 
upper row, and that point being above the tuyeres where the 
heat is greatest will burn out and cause great trouble and 
unnecessary expense, to say nothing of filling the furnace 
with water and freezing it up. The writer has had experience 
with furnaces built in this way and is thoroughly convinced 
that no greater mistake can be made than putting one set of 
jackets above another. 



SPOUTS, SETTLERS, AND JACKETS. 35 

The height necessary for the jackets to extend above the 
tuyeres is dependent upon how the furnace is to be run. If 
it is to be run on high blast and crowded to its greatest 
capacity, then the jackets should extend to the top of the 
charge, say 10 feet above the tuyere centers, but if there is 
to be low blast and easy running, as in lead smelting, then 
they do not need to be so high, since a shaft of firebrick will 
serve quite as well. Generally speaking, the jackets for a 
copper furnace should extend from 8 to 10 feet above tuyere 
centers, which should be from 18 to 24 inches above the bot- 
tom, making the jackets from 9| to 12 feet in total height, 
and of a width depending on the number used and the size of 
the furnace. 

The jackets for the Anaconda furnaces were made of about 
the same height as those for lead furnaces. This was found 
to be a mistake, as the burning out of the shaft proved later, 
and to correct it a series of coils of 2^-inch pipe were put in 
as a lining to the shaft of the furnace from the top of the 
jackets as far up as the brick burned out. The pipes were 
first put together with cast ells, but these broke and malle- 
able ells had to be substituted. 

There were three pipes to the foot and about six feet of the 
shaft were lined, so that it required about 500 feet of pipe and 
54 malleable ells. The water was brought in at the bottom coil 
and made the circuit of the furnace eighteen times before 
escaping. This method of protecting the shaft is much 
cheaper than high jackets, and in case the water does not 
scale or contain much mud, will give equally good results. 
The method is mentioned here for the purpose of showing a 
way of changing a lead furnace to a copper furnace. 

Under stress of circumstances, and in a locality whither it 
would be difficult to transport large jackets, a blast furnace 
could be easily and cheaply constructed with a jacket of pipe 
2| to 4 inches in diameter on the same lines as the brickwork 
was protected at Anaconda. Starting from the crucible or 
base-plate the coil should be put in place, each turn slightly 
larger than the one below to give the shaft the desired taper 
of about one inch to the foot. When the level of the tuyeres 
is reached a nipple of 4 inches in length should be put in and 
the next coil raised by that amount, leaving a space 4 inches 



36 LEAD AXD COPPER SMELTING. 

wide all around the furnace, which should be bricked up at 
all points except where the tuyeres are to enter. Above the 
tuyeres the coils should be extended to the desired height and 
clamped securely together at the corners by means of stirrup 
clamps around the pipe, extending through an iron bar and 
secured by nuts on the clamps. 

The number of coils, malleable ells, and lengths of pipe 
to be used depend on the size of the furnace, the diameter of 
the pipe, and the height the jacket is to be carried. For a 
furnace 40 by 100 inches and a ] 0-foot jacket of J:-inch pipe, it 
would require approximately 600 feet of pipe and 100 ells. This 
would make a jacket that would do just as good work as the 
more expensive steel jackets of large size and great weight. 
The water should be brought in at the bottom and forced 
through the entire coil under a pressure of not less than 25 
pounds. If this pressure is not obtainable then the coil 
should be made in two sections, one above the other and 
supplied from a 6-inch main. The jacket made in this way 
will not be open to the same objection as if one steel jacket 
were placed above another, on account of the flushing out of 
any sediment by the rapidly flowing water. 

Many kinds of fixed tuyeres have been introduced of late 
years to replace the tuyere bags, but they are not an improve- 
ment and are open to many objections to which the simple 
canvas tuyere bag is not. If, for instance, a jacket is to be 
removed they are difficult to manage, and, on the other hand, 
a tuyere bag can simply be twisted and turned out of the 
way. In case slag comes down from above and fills a 
stationary tuyere, it is difficult to remove and frequently the 
tuyere is broken in cleaning. But if bags are used the tuyere 
can be removed and the opening plugged temporarily with 
clay. Good heavy duck bags, well soaked in mineral painty 
with tuyere points of cast iron, are preferable to all the patent 
discharge, slag-catching devices that have been introduced. 



CHAPTER YI. 

Blowing-ik and Barring-Down a Furnace. 

The method of blowing-in a lead furnace, commonly in use, 
is to fill the crucible with molten lead by first firing the 
furnace with wood for several hours to insure its being hot, 
and then throwing lead on top of the wood fire. A blast 
of air is blown down through the lead well or from a tuyere 
in front or back of the furnace to force the fire. By this 
means 250 bars of lead, or about the amount required to fill 
the average-size crucible, can be smelted in ten hours. This 
is generally done on the nightshift, and the furnace cleaned 
of wood ashes by 7 A.M. A layer of fine wood or charcoal is 
then put on top of the lead to a height of a foot above the 
tuyeres, a fire started and the tap- jacket put in place. The 
tuyeres are left open to admit air to the fuel, and after it has 
been thoroughly lighted and fire shows in all the tuyeres the 
furnace is filled up to the top of the jackets with coke. Slag 
charges are then put on alternating with about 12 per 
cent, fuel, and after four or more slag charges alone then one 
slag charge to one ore charge, or two slag to one ore, accord- 
ing to the conditions of charges, etc., until the furnace is filled 
up to about 10 feet above the tuyeres. A light blast is 
started and gradually increased to about one-half the regular 
amount when slag is first tapped. From this time the blast 
is gradually increased for about four to six hours, when the 
furnace should be running on full blast, usually at a pressure 
of about 16 inches of water, but this pressure is dependent 
upon the fineness of the charge and has to be varied according 
to the conditions governing the special case. 

The barring-down of lead furnaces to prevent the formation 
of blowholes and consequent high lead loss, is made necessary 

(37) 



38 COPPER AND LEAD SMELTING. 

by the increase of sulphide and zinky ores treated within 
recent years. In case the furnace is of the stack type with side 
feed-doors this can be done by the regular furnace crew in 
about one shift more or less, depending on the condition of the 
furnace and the energy and ability of the men. The furnace 
should be run down by the nightshift so that the breast- 
jacket can be removed the first thing in the morning. The 
crust is then broken in and the charge raked out and removed. 
Two sets of men then start to cut out the shaft from above, 
while another set work below removing the crusts as they 
drop down the shaft. 

After cutting down the wall accretions from above and 
removing all loose material from the furnace, a large hole is 
cut in the crust over the lead well and the furnace is ready to 
resume operations again with as good results as if just blown 
in. The wall accretions, if allowed to increase in thickness, 
will force the passage of the blast through a constantly decreas- 
ing space, resulting in greater losses both by flue-dust and 
volatilization. In case the furnace is of the top-feed type the 
work of barring down is much greater on account of the greater 
distance the workmen are from the crusts to be removed, 
and in many cases it is found better to blow out such furnaces 
than to bar down, for the reason that while the crusts can be 
removed in the shaft as far down as the top of the jackets, to 
work below there with bars 20 to 22 feet long is exceedingly 
slow and difficult. 

Barring-down in some cases is only partial and is preceded 
by charging the furnace with coke to serve as a bed foi' the 
barrings to fall on. After barring in this way, which is gen- 
erally practiced in the top-feed furnaces, the furnace is filled 
up with charges and operations are resumed. But it is not by 
any means so satisfactory as if carried on until the shaft is 
entirely clean down to the tuyeres, as is the case with side- 
feed furnaces, and is generally not done more than once in a 
campaign. In case the furnace is barred down completely, as 
first described, it is started in much the same wa}^ as if it were 
to be blown in afresh. A wood fire with coke on top to the 
depth of two feet is made, then from five to ten bars of lead 
are added, according to requirements, to fill up the hole cut in 
the crucible, next a few slag charges and then the regular 



BLOWING-IN AND BARRING-DOWN A FURNACE. 39 

charge to the depth of about six feet, when light blast is 
turned on and gradually increased as the furnace is filled up. 
To run a furnace down in the best way, either for the pur- 
pose of blowing-out or barring-down, it is necessary to reduce 
the blast, discontinue charges, and have on hand a consid- 
erable quantity of coke fines thoroughly wet, to be thrown in 
a little at a time as the fire becomes too hot. In the case of 
a copper furnace which is running into a f orehearth, it should 
be allowed to discharge into the forehearth until such time as 
the stream from the spout is becoming too small to continue 
to run, when a bar should be quickly driven in the front and 
the furnace tapped on the side and allowed to discharge the 
remainder of the slag and matte until the blast is taken off. 
In cases of shutdown for any length of time the forehearth 
should be tapped and a bar driven in the matte-tap. 

In the case of a lead furnace it results in too great loss of 
lead to run down to the tuyeres, although there may be four 
or five feet of wet coke fines on top ; therefore the blast 
should be taken off about the time the charge is down to the 
top of the jackets and the remainder of the charge raked out 
on the floor. 

It frequently happens from various causes that a crust will 
form over the lead in the crucible and force the lead out with 
the slag and matte. A sharp lookout has to be kept on the 
slag-pots at all times to prevent this, and, in cases where it 
is possible, to remedy it at once. The causes maybe low lead 
charge, small matte production, bad reduction in the furnace, 
bad slag, water leaking into the furnace, or any one of a 
number of causes too numerous to mention. The quickest 
way is to drive a bar through the taphole down into the 
crucible, but if the condition of the furnace is not changed by 
removing the cause, it will soon become impossible to keep 
the hole open by mechanical means. It is very seldom that 
conditions of charge and furnace are such that it is at no 
time necessary to drive a bar into the crucible. 

It is stated frequently by men who pose as disciples of 
infallibility, that "you should not drive bars into the cru- 
cible." It is also worthy of comment that furnaces run by 
these same men have used up considerable steel and have 
developed some very good strikers. While not wishing to 



40 LEAD AND COPPER SMELTING. 

bear with unusual emphasis on this point as a mark of success, 
or to state that it is necessary for a metallurgist to help his men 
by doing manual labor, still it is advisable in cases of this 
kind either to take hold with both hands or leave it entirely 
to the foremen and men. 

The blowing-in of a copper furnace on a matte charge is not 
attended with as much difficulty as is the case with a lead fur- 
nace. It is important in the starting of any kind of a fur- 
nace that the water connections should all be looked after 
before the furnace is started, in order to provide for the in- 
creased use of water during the short period of time before 
the jackets have become coated with a protecting layer of 
chilled slag. Before blowing-in a matte furnace the hearth 
as well as the settler should be well heated with a wood fire 
for several hours, the spouts being thoroughly dried to prevent 
explosions by contact with matte. 

After this it is only necessary to increase the wood fire in 
the furnace sufficiently to insure a thorough lighting of the 
coke covering, which should be put on to a depth of about 18 
inches above the tuyeres. When the mass of coke is on fire 
throughout, charges of one-half matte and one-half impure 
slag should be put on with coke until the furnace is full. The 
blast is turned on and gradually increased until the full blast 
is reached. As soon as slag shows at the tuyeres, the bar is 
removed from the taphole and the slag and matte allowed to 
flow continuously into the forehearth. It is better to follow 
this plan of filling up the settler, especially if it be a large one, 
than to start at once on ore charges, because it insures the 
entire mass in the settler being in a fluid condition and avoids 
the possibility of a hard matte-tap. If the matte and slag are 
not to be had for the purpose of blowing in, a charge should be 
figured that will produce the greatest quantity of matte 
together with a slag that shall run as easily as it is possible to 
make out of the ores to be smelted. In such a case the coke-bed 
should be somewhat thicker to insure the ores being smelted 
before arriving at the tuyeres. It is very important in cases 
where the settler is large that the slag and matte should enter 
the settler fast enough to float the crust that will form on top 
of the fluid mass. In order to assist this as much as possible, 
men should be stationed around the settler with bars long 



BLOWING-IN AND BARRING-DOWN A FURNACE. 41 

enough to be able to break the crust in a line next the walls 
of the settler, thus allowing the crust to float freely as the 
settler fills. The slag-tap should be closed to force the settler 
to fill up about six inches above its usual height when run- 
ning. This is done to raise the slag-crust into its proper posi- 
tion before allowing it to cool. 

When it has risen to the proper height the slag-tap should 
be opened very carefully and only enough slag allowed to 
escape to keep the crust from rising still higher, until such 
time as it has chilled to a sufficient thickness to support itself 
when the fluid mass is tapped from under it. It may require 
four to five hours and an occasional sprinkling from a hose to 
allow the crust to become thick enough. Sometime before 
the settler is full the slag and matte charges should be taken 
off gradually and ore charges substituted, so that about the 
time the settler is full the ore charge should be down to the 
smelting zone. 



CHAPTEK YII. 

Handling Blast-Furnace Slag. 

The handling of the slag from blast furnaces is an important 
part in their management, and there are four methods now in 
use for lead furnaces and two for copper which have consider- 
able merit. The first to be considered will be the methods in 
use by the Arkansas Valley Smelting Co. at Leadville, at the 
Omaha & Grant works in Denver, by the Mexican Metallurgi- 
cal Co., San Luis Potosi, Mexico, and the Pueblo Smelting & 
Eefining Co. at Pueblo, Colo. At each of these places one of 
the four methods for lead furnaces is in use and can serve as 
an example. 




Fig. 10.— Matte-pots used at Arkansas VAx,iiEY Works, Leadville. 



First, at the Arkansas Yalley Smelting Company's works 
the slag and matte are tapped from the furnace into pots of 
the ordinary or Devereaux type (Fig. 10), and these are run 
on top of a reverberatory furnace built on a lower level, poured 
or partly tapped through the Devereaux hole into the furnace, 
where matte and slag are kept in a fluid condition until a com- 
plete separation takes place between them. The slag is then 
allowed to flow off into larger pots on cars, which are hauled 
away by an engine, while the matte, when it has accumulated 

(4n 




Fig. II. — Matte Settling Pots at Omaha Grant Works, Denver, Colo. 



HANDLING BLAST-FURNACE SLAG. 



43 



sufficiently, is tapped from the furnace into beds. In this 
way the settling is an entirely distinct operation from the 
smelting, and one forehearth serves for several furnaces in 
blast. 




Fig. 12.— Matte settling pots used at Omaha & Grant Works, Denver, 
Colo. ; LoNGiTUDiNAii section. 



The second method to be considered is in use at the Omaha 
& Grant smeltery in Denver (Figs. 11 to 13), where the slag is 
tapped from Devereaux pots into very large cast-steel pots of 
the same shape as the Devereaux with the hole about two feet 
from the bottom, and another hole in the bottom for a matte- 
tap. These large pots are about five feet in diameter at the 
cop, conical in shape, and about six feet deep. They are 
handled by a crane and traveler, of very simple but effective 
design, and can be lifted about and changed with ease. When 
receiving slag from the Devereaux pots they are sunk in a pit 
in the dump so that the top of the pot is about an inch above 
the iron plates which surround it and serve as a floor for the 
slag-pots from the furnaces. Several small pots can be tapped 



44 



LEAP AND COPPER SMELTING. 



into the large one at the same time, and there being two large 
pots in use one can be drawn off while the other is filling. 
There is also a dumping pot on a car pulled by a horse to 
take away the slag after settling the second time (Figs. 14 
and 15). The plan of operations is to allow the slag to settle 
as much as it will in the slag-pots, then to tap out the 
upper portion into the large settling-pots and allow it to 




Fig. 13.— Matte settling pots usei> at Omaha & Grant Works, 
Denver, Colo.; cross-section. 

settle again. When the large pot is full of slag it is tapped 
(as it stands in the pit) into the dump-pot on the car before 
mentioned, and this slag is hauled away to the face of the 
dump. If the large settler has not enough matte in it to 
necessitate a change, it remains where it is and more slag 
is tapped into it after closing the Devereaux hole, and by 
repeatedly filling and emptying it of slag the matte that 



HANDLING BLAST-FURNACE SLAGS. 



45 



escapes from the small slag-pots accumulates until it is 
necessary to change the large settler. It is then hoisted and 
carried by the overhead traveler to a place provided, where 
it can be tapped from the bottom and the matte allowed to 
escape into smaller pots or beds. Meanwhile another large 
settler is put in its place and the process goes on. The object 
of this second settling is to collect and recover that portion 
of the matte which at times will unavoidably escape from the 
tapping of the Devereaux pots. 




Fig. 14.— SliAG-TBUCK USED AT OMAHA & GRANT WORKS, DENVER, COLO. 

The third plan, and the one practiced at San Luis Potosi 
Mexico, by the Mexican Metallurgical Co., necessitates the 
use of overflow pots at the furnace, of a size sufficient to 
collect the matte for from three to four hours, the overflow 
slag going into the Devereaux pots which are again tapped 
into other slag-pots of the same size, the contents of the latter 
being thrown over the dump. The overflow pots are emptied 
into cast-iron moulds arranged in a circle so that a jib crane 
can be used to lift out the matte cakes when cool. 

The fourth method is in use at the Pueblo Smelting & 
Refining Co.'s works at Pueblo, and also at the East Helena 
works at Helena, Mont., which makes the lead furnace a 
constant discharge, and the matte is tapped at a lower level 
than the slag, thus making a settler of the furnace. 

At each of these places the visitor is given to understand 
that the method in use is better than all the others, and it 



46 



LEAD AND COPPER SMELTING. 



is a difficult matter to decide. However, each has points in 
its favor to recommend it, and it is to be supposed that the 
results are about the same. 

In case a copper furnace is to be run with intermittent tap, 
the slag would be handled by one of the methods described 
for lead furnaces. With a constant flow and settler, the only 
two methods in use are, first, by pots, which may be either of 
the ordinary kind on cars and run on tracks, or, second, by 
granulating with water. The first method needs no special 
mention except to state that the slag can be handled more 
cheaply by pots on cars than by the ordinary slag-trucks. 




[(..r ■ '=^' 










Fig. 15.— Slag-truck used at Omaha & Grant Works, Denver, Colo. 



The granulating and sluicing away by water necessitates a 
considerable amount of fall to the sluice, about 5 per cent, 
being the minimum, and about twice as much water is required 
as the jackets will use. A six-inch pipe with a pressure of 12 
pounds will furnish enough water for the jackets and granu- 
lating slag for two 40 by 100-inch furnaces running 125 tons 
each per day, the water discharged from the jackets being run 
into the slag sluice. 



CHAPTER VIII. 
Design of Lead Blast Furnaces. 

The tendency of recent years in constructing lead furnaces 
has been to increase the height from tuyere level to feed floor 
as well as to increase the area of the furnace. 

The dimensions of a lead furnace at the tuyeres have no 
particular bearing on anything except the capacity, but the 
height of the furnace above the tuyeres has a very decided 
bearing on the lead losses, as the experience with the latest 
type at Aguas Calientes will show. A simple statement of 
the facts as they occurred will be necessary to show how a lead 
furnace should not be constructed. 

The plant at Aguas Calientes originally consisted of a con- 
centrator of about 120 tons capacity, two roasters of a type 
that no one need desire to imitate, two copper blast furnaces, 
42 by 120, with forehearths 10 feet in diameter, and a copper 
converter 8 feet in diameter by 16 feet high. It was decided 
to add to this by building two lead furnaces. The copper 
furnaces were built up on a high pedestal of masonry so that 
a ladle could be run below the settlers in a position to catch 
the charge of matte when tapped for the converters. This 
made the foundation for the lead furnaces considerably lower 
than the copper furnaces. The charges for both sets of fur- 
naces had to be elevated to the feed floor, and it was im- 
portant that the charge floors should be on the same level so 
that in case either elevator was under repair the other could 
be used for all the furnaces. The lead furnaces were accord- 
ingly run up until the charge floor was on a level with that of 
the copper furnaces. This made them 27 feet high from the 
furnace floor to the charge floor, or about 5 feet higher than 
top-feed furnaces are built in Colorado. There was nothing 

- (47) 



48 LEAD AND COPPER SMELTING. 

unusual about the furnaces except their height and the feeding 
device, which was in imitation of the bell and hopper used 
with iron blast furnaces. 

On December 26, 1895, one of these furnaces was blown in 
on a charge containing 14 per cent, lead and a slag of about 
33 per cent. Si02, 37 per cent. FeO, and 18 per cent. CaO. It 
started off as furnaces generally do, with all tuyeres bright, 
slag hot, and producing lead a little too rapidly on account 
of the displacement of lead in the crucible by slag and matte 
accumulations. 

The furnace was put on full blast about 6 P.M. and ran 
fairly well until about 4 A.M. the next morning, when the 
tuyeres had blackened and overfire started. The lead produc- 
tion ceased entirely by 12 o'clock M., and from that time 
forward until it was shut down the furnace did not produce 
any more lead. The charge was changed many times ; slag and 
matte charges were fed and the blast was reduced in the hope 
of getting the overfire down. The bell and hopper feed was 
perfectly tight, and the furnace was closed up on top as tight 
as a tin box. The furnace continued to make slag and 
smelted a fair tonnage of ore, but produced no lead. The 
writer was of the opinion that the bottom of the crucible had 
sprung a leak and the lead was going into the foundation. 
The furnace was then shut down and the other furnace, 
which had been put in readiness, was blown in. The charge 
used did not differ greatly from that put on the first furnace, 
though the fuel was increased from 12 to 14 per cent, and the 
blast pressure reduced, but in spite of these precautions and 
the most careful attention the furnace got hot on top and the 
lead production stopped entirely in about 24 hours after the 
furnace was started. 

From this point until the furnace was blown out four days 
later, all the changes that could be suggested were tried. 
The slag was made acid and then it was made basic. It was 
run high in lime and low in lime, but no lead came out of the 
furnaces. Bullion was fed back again and it seemed to 
disappear as completely as if it had never existed. 

The lead gradually fell in the well and bullion was melted 
in the pot and poured into the well until it was full, when it 
would gradually disappear again, showing either that there 



DESIGX OF LEAD BLAST FURXACES. 49 

was no lead going into the crucible or that the crucible was 
leaking. After a council of war it was decided to blow the 
furnace out, as too much money was being lost to allow this 
state of affairs to continue. 

Accordingly it was decided that another man should try to 
run them, and Dr. Charles Harbordt was sent there to do the 
metallurgical work. The doctor is a yalued friend of the 
writer, and was offered eyery assistance that the light of i^ast 
experience could giye as to the running of these furnaces. 
One of them was blown in, and to describe its working would 
only be a repetition of what has been said of the other two 
attempts, except that it ran a week instead of four days 
without producing any lead. The furnace was then blown 
out and a message sent to headquarters that no more furnaces 
would be blown in until the general superintendent came 
down. When he arriyed on the scene he blew in one of the 
furnaces, which jDroduced lead for three days, but in a rapidly 
decreasing proportion to the amount on the charge. 

The second furnace was blown in and acted the same as in 
all preyious campaigns, producing a little lead for two days 
and then the lead production became nil. The tuyeres would 
become black and hard, and frequently raw ore could be found 
at the crucible. Shutting off the blast from the tuyeres 
would haye the effect of burning off the nose after seyeral 
hours, but it would only take them as many minutes to 
become black again when the blast was turned on once more. 
Besides, the furnaces were not being run to keep the tuyeres 
bright but to produce lead, and as to this point there could be 
no question — they were a failure. 

The bell and hopper feed was withdrawn from the first 
furnace, and it was again blown in, all the ores and coke being 
thoroughly soaked in water. Frequently the hose was used 
with good effect on the feed floor and gradually the zone of 
fusion was brought down to the tuyeres. The lead production, 
which up to this i)oint had been nil, gradually rose until about 
60 per cent, of the lead on the charge was produced as bullion. 
This was the best that could be done. When the bell and 
hopper feed was again replaced the lead production ceased as 
quickly as if there had been no lead contents in the charge. 
To a metallurgist who had not seen these things as recorded 



50 LEAD AND COPPER SMELTING. 

this might seem to be an exaggerated statement. It would 
appear to be impossible to lose all the lead in the charge if 
so much as 14 per cent, were used. And that is exactly the 
way it looked to the writer when the first two furnaces were 
blown in, but after seeing three more in blast and under the 
supervision of men who have had long experience in lead 
smelting, he became convinced that science was indebted to 
the designer for finding out two things : First, that a lead 
furnace closed in on top will soon show overfire, and the lead 
will be wholly or partially lost as volatilized fume ; second, 
that increasing the distance between tuyere level and feed 
floor beyond the proper height of the smelting column has the 
same effect as closing the top of the furnace. It assists in 
volatilizing lead by preventing the cold air which is drawn in 
above the charge from cooling the top layers of charge. 

After making many attempts to correct the losses it finally 
became necessary to tear out the crucibles, cut oif the shaft 
of the furnace to about the standard height, 22 to 23 feet 
from the furnace floor to charge floor, throw out the automatic 
feeding device, and return to feeding with a shovel. 

It is certainly of great value to the profession to know what 
cannot as well as what can be done, and while this informa- 
tion is all of a negative character, still in many respects such 
an example has its uses. While it was certainly very annoying 
and perplexing at the time, it is all clear now, and the writer 
is very glad to be able to impart the experience to others 
even at the risk of being suspected as a party to the error. 

After blowing out, a sample of the bullion remaining in 
the crucibles was taken, and it was found to assay about 400 
ounces Ag per ton, while according to the charge calculation 
it should only have had 180 ounces Ag, thus showing that the 
lead had been volatilized in much greater proportion than the 
silver, and that there had been a cupelling action in the 
furnace. 

After a careful sampling and weighing of all products and 
giving credit for the lead in the slag produced, which was all 
high in silver and lead, it was found that 63 per cent, of the 
lead and 23 per cent, of the silver were unaccounted for. 

The deductions that can be drawn from this failure are 
very simple and exceedingly important. They point clearly 



DESIGN OF LEAD BLAST FURNACES. 51 

and with force to the fact that the first lead furnaces in use 
in this country were better adapted to the saving of lead 
than the modern high-shaft, top-feed furnace. The furnaces 
with the charge doors on the side, a stack above the floor, 
with downtake from this stack to the flue beneath the floor, 
as originally constructed in Leadville and at the Colorado 
Smelting Co.'s works in Pueblo and also at Great Falls, 
Mont., nearly fulfill the opposite of all conditions that existed 
at the furnaces at Aguas Calientes. In the first place, the 
amount of flue dust and the loss resulting therefrom is much 
less in a stack furnace than in a top feed, for the very suffi- 
cient reason that the flue dust must rise a distance of 10 to 14 
feet before entering the downtake of a stack furnace, and, on 
the other hand, in a top feed the downtake is right at the top 
of the charge, where every inducement is offered for fines to 
go into the flue and thus heavily increase the mechanical loss. 

In a stack furnace the air entering the charge doors comes 
into contact with the top of charge, thereby preventing con- 
siderable lead losses. That the cooling of the top of the 
charge in the furnace by the air drawn in at the charge doors 
has this effect is proved by the excessive losses in the furnaces 
at Aguas Calientes, where the only point of difference was 
bad draught and excessive height of shaft above the top of the 
charge, which prevented the cold air having access to the 
charge. The downtake was immediately below the charge 
floor, and the furnaces were fed at various depths from the 
downtake to thirteen feet below the charge floor. They pro- 
duced no lead when the smelting column was 18 feet above 
the tuyeres, and did the best work when it was 10 to 11. 

This experience, disastrous though it was, was one of the 
greatest object lessons that has ever been given to lead 
smelters. It shows plainly that lead has to be treated as a 
very volatile metal closely resembling mercury in its behavior, 
and that if the greatest possible saving is to be made the 
blast pressure must be light, the smelting column not much 
over 12 feet above the tuyeres, and the greatest possible 
amount of cold air must be allowed to enter the furnace at 
the top of the charge. If these points are well taken, then all 
the top-feed thimble furnaces that have been built so exten- 
sively in the last ten years are steps in the wrong direction. 



52 LEAD AND COPPER SMELTING. 

There is another point about the stack furnace that is much 
in its favor, and that is, being fully ^ve feet shorter from the 
charge floor to the furnace floor, it is much easier to bar down 
when accretions form and, by doing this once a month, can be 
kept in blast for any length of time desired. 

No doubt there will be many to disagree with these conclu- 
sions, and the writer will admit that he was of different mind 
until the force of experience brought out the points in 
question. 

Briefly stated, the points of superiority of the stack furnace 
over the top feed are : 

1st, It makes less flue dusto 

2nd, It runs cooler on top. 

3rd, It loses less by volatilization. 

4th, It is more easily barred down. 

5th, It can be kept in blast longer. 

At the Arkansas Yalley works the furnaces as originally 
built were of the stack type. They were designed by Mr. Eilers, 
and the same type of furnaces are now in use at all the works 
where he has had charge. It has come to be a matter of 
comment that a man posted in the hobbies of metallurgists 
can tell by the appearance of a plant who was its designer. 
Certainly in this case the hobby was of the right sort. 

Many kinds of thimbles and charging devices have been 
introduced and used on lead as well as on copper furnaces, and 
they have as often been discarded, the method of feeding 
with the shovel into an open-top furnace thus far being found 
superior to mechanical feeding for many reasons. 

First, the furnace does not run alike in all parts and 
requires to be fed the greatest amount at the point where it 
sinks the most rapidly. Second, certain kinds of material 
must be put in particular places in order to correct irregular- 
ities of running. Third, any approach towards closing in the 
top always has the effect of drawing the fire up, resulting in 
increased losses by volatilization. Fourth, this loss by vola- 
tilization in case of lead will increase rapidly as the air is 
prevented from entering at the top of charge. 

The best method of feeding is to have a fender in front of 
the charge door (to prevent as much as possible the dumping 
of charges into the furnace either accidentally or intentionally, 



DESIGN OF LEAD BLAST FURNACE. 63 

but more often the latter) and to scatter the charge thor- 
oughly over the surface of the coke which has been shoveled 
in in the same way, but with more on the sides near the 
furnace walls than in the center. The reason for putting the 
coke next the furnace walls is that as it goes down it may 
burn off accretions, and when it has arrived at the tuyeres 
the heat will be driven into the ore. Sometimes accretions of 
considerable size can be taken from the walls of the furnace 
by persistently feeding the fuel against them. From the very 
nature of things it will be apparent that a chance distribution 
of the charge, as would be the case with bell and hopper, is 
the poorest possible contrivance for feeding a lead furnace. 
It worked on the copper furnaces at Aguas Calientes, though 
very unsatisfactorily — principally because it was impossible 
to keep the furnace cool on top, the result being high silver 
losses and the consumption of the fuel before it arrived at 
the proper smelting zone. 



CHAPTER IX. 

Lead Slags and Losses in Lead Smelting. 

The calculation of slags for lead furnaces is carried on in 
exactly the same way as for copper work, Avitli the additional 
care that has to be taken to keep a sharp lookout for the zinc, 
sulphur, and baryta. The ores have to be bedded in such a 
way as to admit of their being used to the best advantage, and 
this is an absolutely necessary precaution to insure success. 
Frequently it may be of advantage to change the charge, to 
put on more or less of some class of ore, either lead, dry or 
sulphide, and unless the ores are bedded according to these 
classifications it might be difficult to make the required 
alteration. 

The following is a fair example of lead charge as made up 
from beds : 







6 

m 


O 


d 


O 

0) 


d 


i 


S 




a: 










•Jl 


^ 


fe 


^ 


o 


^ 


f^ 


_j 


■c 




-u 




m 




m 


rH 


m 


r- 


■j: 


— 


CQ 




^ 




^3 




■tJ 


c:^ 


'd 


^ 


"C 




-a 




sc 


o 


fl 


o 


C 


o 


C 


U 


C 


o 


a 




'S 


u 


S 


u 


S 


•- 


P 


k. 


5 


%-, 


s 




K 


o 


O 


o 


Q 


o 


o 




3 




o 




> 


Ph 


^ 


u 


"^ 


U 


Ph 


^ 


rH 


cu 


Dh 


Carbonate 


4C)0 


i(^ 


104 


hi 


124 


() 


24 


12 


4S 


4 


hs 


(Silicious lead 


200 
100 


41 
15 


82 
15 


12 
85 


24 
35 


9 




4 



10 
31 


20 
31 


6 
5 


12 
5 


Roasted 


Iron ore 


100 


10 ■ 


10 


55 


55 


2 


9 


2 


2 






Ijime 


200 


3 


6 






52 


104 


















1000 




217 




238 




134 


. . 101 


• • 


33 



Owing to the presence of some lead and zinc in lead slags, 
the sum of the Si02, FeO, and CaO do not, as a rule, amount 
to more than 88 per cent., so that figure has been used in this 
calculation. Also, it has been assumed that 20 pounds of the 
33 pounds of sulphur on the charge will go into the matte 
while the other thirteen will be burned ofi" as sulphurous 

(54) 



LEAD SLAGS AND LOSSES IN LEAD SMELTING. 55 

acid. Twenty pounds of sulphur would indicate about 80 
pounds of matte, as lead mattes carry about 25 per cent. S, and 
the iron equivalent of 40 FeO, so that 40 per cent, of 80 pounds 
equals 32 FeO, and that amount should go into the matte, 
leaving 206 pounds of FeO for the slag. 

This type of slag, 34 SiOs, 33 FeO, 23 CaO, is a favorite 
with many smelters, and is largely used in the smelting of 
copper as well as of lead, although in copper work much less 
lime will do just as well and result in treating more ore. 
With only 33 pounds of sulphur on the charge the matte 
formation would in all probability not be as much as 80 
pounds, and very likely not more than 50 or 60 ; if the lower 
figure should prove to be correct, then only 20 pounds of FeO 
should be deducted, which would change the resulting slag 
composition to 33.4 per cent. Si02, 34.1 per cent. FeO, and 21 
per cent. CaO, which is just as good a slag as the first, and 
whichever is used, there would be no material difference in 
cleanness of work. Slags as low as 30 per cent. Si02, and as 
high as 40 per cent. FeO and 20 per cent. CaO, are perfectly 
safe, and formerly, under certain peculiar conditions which 
now unfortunately do not exist, were good commercially. 

Generally speaking, iron slags are more fusible than lime, 
and as a consequence give rise to less loss in the volatilization 
of lead and silver, and they will absorb more zinc without 
becoming dangerously hard in the taphole. 

The peculiar feature of high lime slags is their remarkable 
flint-like toughness when they once get cold. A furnace 
running on such a slag may be perfectly free and all right, 
and owing to a hard breast may have slag running out of half 
the tuyeres fifteen minutes later. 

The writer has in mind one superintendent who if he has 
one particular hobby more than another it is to run on high 
lime. No matter what is to be accomplished, it can only be 
done by raising the lime. If the reduction is to be improved, 
the fuel cut down, the furnaces run faster, or what not, just 
raise the lime and, presto, you have it. Another peculiarity 
with certain persons is that by some optical method of 
analysis the lime is always higher than a chemist can get by 
any known method except the " graphite." 

It is certainly a bad plan for a metallurgist to insist on 



56 LEAD AND COPPER SMELTIXG. 

having the slag anah'ses agree with his calculations, and to 
insinuate that the chemist does not know what he is about 
simply because he reports the results as he gets them. It 
might appear that this is totally superfluous, but the writer 
has lived in the same atmosphere with a great many men who 
would forget all the possibilities of mistakes that could be 
made in weighing charges, in bedding ores, of throwing on 
the wrong ore at the scales and a shovelful more or less, as 
the fancy struck them, and then blame the chemist for in- 
accuracy of work. Besides, slag as it comes from the furnace 
is not of a constant composition. Every pot is slightly 
different from every other pot, and a close approximation is 
all that can be expected. 

The losses in lead smelting are generally estimated at 5 per 
cent. Ag and 10 per cent. Pb, and in making purchases of ore 
it is customary to make this deduction from the metallic con- 
tents of the ore in making payment. The actual losses, or such 
as are shown in statements, vary according to local conditions 
of charge, plant, and ability of the metallurgist. 

With three unknown quantities in the equation, each of 
which is subject to great variation, it is not surprising that 
the known results should vary all the way from a 5 per cent, 
gain of silver to a 23 per cent, loss, and from a 2 per cent, 
gain of lead to a 63 per cent. loss. The latter figures for high 
losses cover the results of the different campaigns at Aguas 
Calientes, with the bell and hopper feed, and the gains, which 
were equally remarkable, were obtained at the Arkansas Val- 
ley Smelting Co.'s works, Leadville, by the writer while treat- 
ing a pile of waste material, the contents of which were not 
charged on the books. It will, of course, be understood that it 
is impossible to make a gain unless some such conditions as 
those mentioned are present, where the recovery of something 
supposed to be lost is counted as a gain. 

Properly speaking, the losses in lead smelting can be kept 
within the allowed limits of 5 per cent. Ag and 10 per cent. 
Pb, but to do this it is necessary that the furnaces shall not 
be run on too low a lead charge, say not below 11 per cent., or 
the tonnage crowded too much by high blast. 

The losses of lead in the slag will increase as the amount 
on the char2:e decreases, for the reason that there will be more 



LEAD SLAGS AND LOSSES IN LEAD SMELTING. 57 

slag, and with the same contents of lead it would represent 
a greater percentage of loss. In addition as the lead on the 
charge decreases the charge becomes more infusible, and as 
a consequence a greater amount of lead is lost by volatiliza- 
tion. 

The writer has had returns from months during which the 
results on accurate charging of all contents represented a loss 
of 4 per cent. Ag and 8 per cent. Pb, and in other months 7 
per cent. Ag and 1-i per cent. Pb. At all of the Colorado 
smelteries there is an effort to crowd the tonnage of the fur- 
nace as much as possible, and this can only result in heavy 
losses of silver and lead. There is also considerable roasting 
and slagging of sulphides as a preliminary step to smelting, 
and it is safe to say that 5 per cento of the silver and 15 per 
cent, of the lead contents are often lost in this one operation 
alone. If the ores are simply roasted the loss will be very 
much lighter, probably not over 2 per cent. Ag and 4 per cent. 
Pb, but the slagging or fusing of the roasted charges in the 
fuse-box, as practiced at many of these works, is certainly 
productive of very high losses in the general return. Accurate 
returns from slagging a lot of 60 tons of flue dust at the 
Arkansas Yalley smelting works, in Leadville, showed 11 per 
cent, silver and 25 per cent, lead loss. Similar tests on slagging 
ores containing no lead showed 5 to 8 per cent, silver loss in 
roasting and slagging, while the simple roasting loss was below 
1 per cent. If any considerable portion of the tonnage of a 
smelter is treated in this way, the smelting losses cannot be 
expected to be within 5 and 10 per cent, limits. There is 
another feature that plays an important part in smelting 
losses and that is the amount of by-products or matte, 
barrings, and flue dust produced. If the contents of these by- 
products are credited to the smelting account at any more than 
95 per cent, of the full amount, an error is made, for they 
have to be smelted again, and it is safe to say that in doing 
so 5 per cent, of the silver and 10 per cent, of the lead con- 
tents will be lost. 

The percentage of fuel necessary to do good work or to 
smelt different classes of ore at different plants is subject 
to great variation. The simple fact that a furnace can be run 
-^n 9 to 10 per cent, fuel basis is not proof sufficient that the 



68 LEAD AND COPPER SMELTING. 

addition of 2 or 4 ]3er cent, would not smelt enough additional 
ore to make it a better commercial proposition. For examxDle, 
at Denver, Pueblo, or Leadville, with coke at $6 and the 
average charge margin at $4, a furnace that would smelt 40 
tons a day on a 10 per cent, fuel charge, would probably 
smelt 45 tons on 12 per cent. fuel. In the first case it would 
use 4 tons coke, $24 ; margin, $160. In the second it would 
use 5.10 tons coke, $30.60; margin, $180; showing a gain of 
$13.40 per furnace in favor of higher fuel. This is only true 
within certain limits, however, because the increase of fuel 
beyond the limits required for smelting and reduction does 
not increase the tonnage, but rather decreases it owing to 
the extra time required for the combustion of the surplus 
before the charge can come down. 

The effect of different kinds of coke on the running of the 
furnace as well as upon the losses is very marked, although 
the reason is at times very obscure. There is a coke made at 
Sabinas, Mexico, which, according to reports of metallurgists 
who have used it at Monterey and Velardena, causes an 
abnormal loss of lead in the blast furnace, owing probably to 
high percentage of ash and volatile matter ; the volatile matter 
IDassing off as gas causes overfire, which in turn volatilizes 
lead. A coke may have as much as 20 per cent, ash and if 
well baked and free from volatile matter may yet x>roduce 
fair results in the blast furnace, although a coke with less ash 
will do proportionately better work. 

It is also customary to increase the fuel accordingly as the 
altitude of the place is higher. At Leadville the fuel con- 
sumption is from 3 to 5 per cent, more than at Pueblo, 5000 
feet lower. The reason for this is that a greater quantity of 
air has to be blown through the furnace to get the required 
amount of oxygen, and the surplus of inert nitrogen produces 
a chilling effect on the furnace, which has to be remedied by 
the addition of more fuel. 



CHAPTER X. 

Improvements in Roasting Furnaces. 

In the past five years many improvements in roasting 
furnaces have been made having for their object the reduc- 
tion of labor and fuel expense in desulphurizing copper ores 
and concentrates. The attempt has been made with the 
Brown- Allen-O'Harra to utilize the heat from the reverbera- 
tory smelting furnace to do the roasting, but owing to the 
breaking of the chain, as well as the interruption of the draft 
at the doors where the plows pass in and out, it has been 
found impossible to keep the furnace in constant operation 
or to develop the amount of heat required. To overcome 
these defects ana to make it possible to roast, smelt, and 
convert with only one firing and with the minimum amount 
of labor, the writer has designed a roasting furnace that can 
be coux^led on to a reverberatory and the heat from the 
smelting utilized for roasting as well as the converting of the 
matte. The great reduction in the price of steel rails in the 
last feAv years is an example of the benefit derived from 
utilizing heat once developed and not allowed to go to waste 
before the finished product is turned out. The cast iron is 
taken direct from the blast furnace to the converter, and the 
converter blooms run hot into the soaking x^it to await the 
rolls, where they are rapidly worked up into steel rails with- 
out having lost the heat imparted by the coke of the smelting 
furnace. If a great economy in the cost of producing copper 
is to be effected much the same policy must be pursued, and 
this design is here offered to show what can be done in this 
direction. (Figs. 16 and 17). 

Of all the mechanical roasting furnaces yet designed the 
O'Harra is the most suitable to be connected with a rever- 
es'.)) 



62 



LEAD AND COPPER SMELTING. 



beratory furnace on account of its long hearth, and the 
fact that it can be made to deliver the calcines into hoppers 
on top of the smelting furnace. The difficulties have been 
with the chain, which was subject to much wear on account 
of passing through the tire, and the opening of the flap-doors 
on the ends of the furnace each time a set of plows passed in 
or out of the furnace. The flap-doors let in so much air 
that the hearth was cooled for several feet near each end and 
the draught of the furnace seriously affected. 




Fig. 18.— The Hixon boasting furnace: cross-section. 



To obviate these drawbacks, and especially to remove the 
chain from the fire, the furnace illustrated m Figs. 18 and 19 
was designed. Two cast-iron conduits are built into the 
hearth of each floor, and the wire rope as well as the wheels 
Avhich carry the plow arms move in these conduits, nothing 
but the plows and the arms being above the slot, the action 
of the plows being similar to that of a planer in a machine 
shop. They move forward and backward on the same deck 
without passing out of the furnace, and plow in only one 
direction. A knot on the rope striking the reversing gear 
throws out the clutch at one end and at the same time throws 
in a clutch at the other end of the furnace, so that the motion 



IMPROVEMENTS IN ROASTING FURNACES. 



63 



is reversed automatically each time the plows reach the end 
of their path. By means of the lever arm shown in the draw- 
ing attached to the rope and turning on the ]3low arm, the 
plows are alternately thrown in and out of action according to 
the direction of their travel. The plows traveling a greater 
distance than they are placed apart, the ore is picked up and 
carried along by one set after another has dropped it until the 
end of the stroke, when the next set of plows comes back of 
the place before taking hold and carries it along to the 
opening through which it falls to the lower hearth. Here the 




Fig. 19 —The Hixon boasting furnace; details of carriage and plows. 

plows act in the opposite direction and the calcines are finally 
delivered into the hoppers over the matting furnace ready to 
smelt. The lower deck is extended on I-beams out over the 
matting furnace, and is covered in so that the ores will con- 
tinue to be as hot as when in the furnace. By the action of 
the draught cold air is drawn in through the conduits and fed 
up into the furnace, at the same time cooling the rope and 
wheels. 

The plows may be removed from the trucks by drawing them 
out through the removable door at the end of the furnace and 
detaching the rope from the lever arm. Another set can be 
substituted and work go forward with but short delay. As 
the plow arm is supported at two points and is drawn by two 
ropes, it is capable of being made much longer than the 
O'Harra plow arm, and the furnace hearth can be built any 
width or length desired, with either single or multiple decks. 

The heat from the smeltino: furnace is taken out of the two 



64 LEi\D AND COPPER SMELTING. 

corner flues to either side of the roasting furnace and led into 
the different roasting hearths on the side, where it passes 
along to the other end and unites again to pass out through 
the iron flues, in which would be deposited the heaviest of 
the dust. The dust from these hoppers can be conveyed b}^ 
a spout to the lower hearth and taken by the plows to the 
smslting furnace. The lateral flues and dampers shown in 
the drawing are provided so that any portion or all of the 
heat from the smelting furnace may be sent through or around 
the roasting furnace to the main dust chamber. This si 
necessary for the proper regulation of the temperature in the 
roasting furnace as well as for emergencies when the plows 
need changing. The slag from the reverberatory can be 
skimmed into a ladle and taken away by car or granulated by 
water, while the matte would be poured from the ladle into a 
converter and blown to copper. 



CHAPTER XL 

Smelting Raw Concentrates with Hot Blast 
AT Anaconda. 

The experiment of smelting raw concentrates with hot blast 
was tried, and while the results were not satisfactory they 
were nevertheless instructive. Two brick stoves, similar in 
construction to iron furnace stoves, were erected by using sec- 
tions of old Bruckner furnaces, which were bolted together to 
form the shell. They were built 45 feet high, 8 feet in diam- 
eter, and were lined with a checker work of firebrick made in 
Anaconda. 

The fuel used was slack coal in Taylor gas producers. The 
gas was burned in one stove while the blast was going through 
the other. The blast could be heated to about 500° F. A fur- 
nace w^as run for some time on a charge of 1000 pounds raw 
concentrates and 1000 pounds converter slag. The matte pro- 
duced contained about 35 per cent. Cu (too low for converting 
at Anaconda), the concentration effected being about three 
into one. 

In the resmelting of the matte with hot blast, together with 
converter slag, the grade was only increased to 45 per cent. 
Cu, showing that after the copper has reached a certain point 
hot blast fails to cause the elimination of iron. About 5 per 
cent, fuel was used during the experiments, but later this was 
increased to eight and hot blast was used in smelting the refuse 
from converting. It effected a saving of fuel as against another 
furnace run on cold blast, until by the alternate expanding 
and contracting the brick lining of the stoves gave out and 
the experiment was abandoned. The saving by the use of 
hot blast was about $25 per day, but the stove linings were 
too expensive to admit of any economy even at that rate. 

. (65) 



CHAPTER XII. 

Copper Converting at Anaconda. 

At the time the writer entered the employ of the company 
for the purpose of constructing a converter plant to treat the 
entire tonnage of matte produced, about 300 tons daily, there 
was in operation a converter plant of twelve vessels that 
had been erected as an experiment. The workings of this 
plant had been very unsatisfactory, and the short life of the 
linings as well as the large losses had caused it to be severely 
criticised. At that time the process of copper converting was 
not very well understood, and it was the vain hope of many to 
find a lining that would last a week, as in the case of steel- 
making. That has proved the great stumbling-block for 
many, though the difference in the processes is apparent. In 
steel-making the charge is blown only long enough to burn 
out the excess of carbon, and before the iron has begun to 
oxidize the converter is turned down and the charge poured 
out. All of the products of combustion in this case are 
gaseous except what Si02 may be formed by the silicon pres- 
ent in the cast iron. This Si02 has a protecting influence on 
the lining of the vessel, since it will combine with any oxide of 
iron formed and supply Si02 that would otherwise have to 
come out of the lining. In copper converting the matte may 
contain from 13 to 35 per cent, of Fe in the form of sul- 
phide, all of which has to be converted into the oxide before 
it can be separated from the copper and sulphur. The 13 
per cent, of iron would indicate a matte of 60 per cent. Cu, 
and the 35 per cent, of iron about 28 per cent. Cu, these being 
the limits within which the writer has converted. 

In the case of a 60 per cent, matte with 13 per cent. Fe, 
each ton of matte charged into the converter would contain 

(60) 



COPPER CONVERTma AT ANACONDA. 67 

260 iDOunds of Fe. A partial analysis of the slag formed after 
the charge has been blown to the skimming point is Si02 37, 
Fe 38, Cu 5, showing that for each pound of iron in the matte 
a pound of silica has to be provided in order to form a slag 
which will, with the limited heat generated by the combustion 
of the Fe to FeO and the partial combustion of the sulphur 
to SO2, remain fluid and admit of being poured out of the 
vessel. It must be borne in mind that it is not only a 
question of oxidizing the iron, but of separating it from the 
charge after it has been oxidized and removing it from the 
converter as quickly as possible. This can only be done by 
forming a fusible slag, and a slag can only be formed by the 
union of an acid and bases, and as the bases are formed 
naturally in the converter the acid must be present to unite 
with them as fast as they are formed, otherwise there will be 
an accumulation of very infusible FeO in the converter which 
will mix up mechanically with the white metal (sulphide of 
copper) and make a spongy, viscid mass which cannot be 
skimmed or finished to copper. Bases do not unite to form 
slags at any practicable temperatures, consequently a basic 
lining is out of the question for copper converting. The oxide 
of iron alone without silica will not form a fluid slag, so that 
if it were possible to run a charge in a water- jacketed con- 
verter without freezing, which it is not, it would still be 
impossible to remove the FeO from the converter because it 
would not separate from the sulphide of copper. 

All chemical unions are attended with the development of 
more or less heat, and it is quite probable that the union of 
FeO with Si02, to form a slag, is attended with a considerable 
heat development, and if the converter were robbed of this by 
the use of any other kind of a lining the heat would be 
insufficient. As for basic lining and water- jacketed converters, 
the writer has tried them both and will give in detail the 
results so that aside from theory the actual outcome may be 
known. 

The first experiment to be tried was to protect the lining 
immediately above the tuyeres where the corrosion is greatest. 
This was attempted by imbedding in the silica and clay 
lining a pipe coil running around three sides of the converter 
and immediately above the tuyeres. This coil was of l^-inch 



68 LEAD AND COPPER SMELTING. 

pipe, put together with malleable ells and clamped to the 
shell of the vessel by stay bolts about 8 inches in length. 
The coil was connected at the inlet by a hose to the water 
main in which the pressure was 45 pounds ; the discharge was 
also provided with a hose connection so that the converter 
could be turned up or down without interfering with either. 
The lining was put in in the usual manner and thoroughly 
rammed behind the pipe coil, and as the lining was about 20 
inches thick at the tuyeres the charge could not come into 
contact with the coil until the 12 inches of clay covering had 
been eaten away. The first and second charges were run 
as usual with the rapid decrease in the thickness of the 
lining and increase in size of the vessel. The third charge 
cut away ail the lining in front of the coil, and when the 
converter was turned down to skim slag the pipes could be 
seen bare and exposed from the nose of the converter. At 
this point the experiment was attended with considerable 
danger, for if the pipe had given way or been attacked by the 
matte an explosion would certainly have occurred that would 
have wrecked the entire plant as well as killed the men 
working on the vessel. For this reason everybody kept at a 
respectful distance while the converter was blowing, and only 
approached when necessary to turn the vessel up or down. 
But nothing having occurred to frighten the men they gradu- 
ally came to regard it as being safe and returned to work as 
usual. 

The charge was finished and the copper poured when it was 
found that large lumps of copper were adhering to the tuyeres 
and pipe, but at other points the lining had been corroded as 
much as usual. A fourth charge was attempted, but the 
bottom of the vessel became too thin and the charge was lost 
on account of breaking through at that point. 

It was thus demonstrated that protecting the lining at one 
point only changed the corrosion to another, so that it became 
a question of abandoning the experiment or putting in pipe 
enough to water-jacket the whole interior. This was done, 
and a coil was constructed and put into the converter which 
should be imbedded in and protect the lining as high as the 
top of the charge. The converter was charged and the first 
charge finished without exposing the pipe ; the second charge 



COPPER CONVERTING AT ANACONDA. 69 

was blown to slag and skimmed but ran cold on the finish 
blow, and more matte was tapped in and blown to slag and 
skimmed again. By this time the pipes were exposed 
throughout the interior of the vessel and the lining had been 
eaten away from behind them in some places. The attempt 
to finish the charge was a failure, and it had to be poured out 
as white metal and used as scrap in the other converters. 
Many other attempts were made with this converter, jacketed 
as it was, but they all ended in failure and demonstrated 
beyond a doubt that a charge could not be blown to slag 
after the pipe became exposed, for the reason that there was 
no silica to flux the basic FeO formed by the oxidation of 
the iron in the matte. Water-jackets were then abandoned 
and a lining composed of burned lime mixed with coal tar 
was tried. The result was a foregone conclusion, but never- 
theless we were hunting for straws to grab at and it was 
given a fair and impartial trial. Just enough tar was used 
to stick the lime together, and when in place the lining was 
baked with a coke fire to drive off the gaseous portion of the 
tar. A charge was run into the vessel and blown until the 
flame at the nose indicated the complete oxidation of the iron. 
The converter was then turned down to skim, but in place of 
fluid slag and matte there was found an indescribable mush 
of matte, lining, and oxide of iron, with which nothing could 
be done, and after one more trial with the same result the 
experiment was abandoned. 

In conclusion, and before leaving the subject, it is well to 
state that the silica lining for a copper converter serves a 
double purpose, that of a lining and also a flux, and that it is 
necessary to the success of the process that the lining should 
be corroded. If this corrosion were stopped by any means 
except the introduction of silica in some other way, the 
process would be defeated. The attempt to introduce silica 
through the tuyeres with the blast has thus far proven a 
failure, and very likely always will, for mechanical reasons. 
In the first place the time during wdiicli SiOa is needed is 
very short, and the quantity required is too large to admit of 
even a small part of it being introduced. 

Take the case of a five-ton charge of matte containing Cu 
35 per cent., Fe 30 per cent., S 26 per cent. Each ton of 



70 LEAD AND COPPER SMELTING. 

matte would contain 600 pounds of Fe, and the five tons would 
require 1,620 pounds of Si02 to be introduced through the 
tuyeres in the space of one hour in order to make a slag with 
25 per cent. Si02, and 60 per cent. FeO. 

The lower grade in copper the matte is, and the higher in 
iron, the more basic will be the resulting slag, and while 55 to 
60 per cent. Ou matte will make slags as high as 40 per cent. 
Si02, a 35 per cent. Cu matte will make a slag with about 25 
per cent. Si02, and 60 per cent. FeO. 

The rapid corrosion has the effect of making the lining lose 
its binding force and it falls off in chunks, just as the rapid 
erosion of the banks of a stream by water will make them 
cave off in large pieces. Consequently the life of a lining is 
not exactly inversely proportionate to the amount of iron in 
the matte, but decreases in length more rapidly than the iron 
increases. For example, a lining that would produce 200 bars 
on 55 per cent. Cu matte with an iron contents of 17 per cent, 
would probably not produce more than 50 on a 28 per cent, 
matte with the Fe 35 per cent. But there are so many chances 
for weak points to develop unexpectedly in the lining with 
the low-grade mattes that no comparison is of much value. 

After testing all the different schemes that could afford a 
possible solution of the difficulties of relining converters in the 
stands, it became apparent that the proposition was a mechan- 
ical one ; to handle the converters as quickly as possible and 
make a change of vessels in the least possible time when a 
lining was destroyed. The practice at the old plant at Ana- 
conda, and for some time at the Parrot, was to run the con- 
verter until the lining became too thin to stand another charge, 
and then to flood the vessel with water from a three-inch hose 
and allow the stream to run until the temperature had been 
reduced so that after the water had been turned out a man 
could go inside, cut out all the loose jagged points and slag 
shell, and then reline the vessel. The cutting out was done 
while the vessel was turned bottom up, and the loose material 
was thoroughly cleaned out before any more lining was put in. 
The converter was then righted and the lining passed in 
with a shovel through the nose in lumps, about 8-inch 
cube. The liner first put in the bottom by throwing the 
lumps of lining against the bottom of the converter and after. 



COPPER CONVERTING AT ANACONDA. 71 

wards pressing them in place with his foot. The bottom 
should be from 4 to 6 inches below the tuyeres and made as 
thick as the side lining. The bottom finished, a circular board 
about the size of a barrel-head was put on top of the clay for 
the liner to stand on, and with this as a form the side lining vras 
built up against the old lining remaining in place, or against 
the shell of the converter if it had all fallen out in the cutting- 
out process. 

Large mitts were used by the liners, and a trowel to cut off 
and shape the lining after the lumps were pounded into posi- 
tion with the fist. Attempts to use all kinds of tamping irons 
were made, but it was found that it took much longer to put 
in a lining with them than with the hands, and there was no 
apparent improvement in its character. 

The lining was made of pure white quartz crushed to the 
size of slack coal in crushers and rolls and afterwards ground 
in a Chilian mill with one shovel of fat, sticky clay to 8, 9, or 
10 shovels of quartz, according as the clay seemed to vary in 
plastic or binding qualities. The clay which was employed 
contained a somewhat large percentage of alkali earths as 
well as iron, and only so much of it was used as was neces- 
sary to stick the quartz together when moistened and ground 

Clay Analysis. 

SiOa 66.0 per cent. 

AI2O3 18.5 

Fe. 3.1 

CaO 2.9 

H2O 8.4 

98.9 

in the Chilian mill. On account of the flooding of the con- 
verter before relining, very little difficulty was experienced 
in making the new lining adhere to the old, but later, when 
the new plant was constructed and the converters were relined 
without filling them with water, it would sometimes happen 
that the fresh lining would part from the old, and linings put 
into a dry vessel did not last as long as those put into one 
that had been made thoroughly wet Notwithstanding this 
small advantage, the use of water for cooling the vessels is a 
poor policy for many reasons, the first of which is that there 
are formed by the blowing of the charge sulphates of iron 
and copper, which are dissolved by the water and react on the 



T2 LEAD AND COPPER SMELTING. 

ironwork of the converter, corroding it rapidly and causing it 
to break open at the riveted seams after a couple of months 
continuous run. The lining remaining in the vessel swells by 
the addition of water and brings a heavy strain on the con- 
verter, frequently causing splits of the seams. Moreover all 
the copper that is dissolved as sulphate is carried away and is 
lost unless recovered on scrap iron. 

At this plant, which, as stated, was an experimental one, 
there were three converters to each cupola or melting furnace, 
and the charge was run from the cupola to the converters in 
long spouts. One of the three converters was kept in opera- 
tion while one was drying out and the third was relining. By 
putting on extra crews when the matte was of good grade it 
was sometimes possible to work five or six of the twelve ves- 
sels at one time. But this was only possible when the copper 
in the niatte was as much as 60 per cent, and the iron as low 
as 13 per cent. Otherwise the linings would be destroyed too 
rapidly to admit of runing more than one vessel out of the 
three. Owing to the bad dust-chamber arrangements and to 
the use of water in the converters, the losses at this plant 
were exceedingly high. During the year and a half that the 
writer ran it, before the new plant could be constructed, the 
losses in copper were 4 per cent, and silver 5 per cent., while 
the cost of converting varied between 0.72 and 1.34 cents per 
pound of copper converted from matte averaging 55 per 
cent, copper. 

The converters were 60 by 60 inches, square in section, and 
turned by a worm gear operated by power from a line shaft. 
The plant was put into a building which it did not fit, or 
rather the building did not fit the plant, and, generally speak- 
ing, everything worked at a disadvantage. The experience 
gained was the basis for constructing the new plant and was 
valuable from that point of view. The largest production in 
any single month from this plant was 5,500,000 pounds of 
copper. 

Previous to constructing the new plant several systems of 
handling converters were developed and their merits dis- 
cussed. It was finally decided that tlie handling by crane was 
preferable to the car, and the detail plans were drawn up 
accordingly. 



COPPER COXYERTIXG AT AXACOXDA. TJ 

It was unfortunate, though unavoidable, that the plant 
had to be built with remelting furnaces, the smelting works 
being already established and so compact in arrangement 
that it would have been impossible to take the matte 
from the reverberatories to the converters in a molten 
condition. This is a point that in any new construction 
should be kept in mind, even though converters are not to 
be installed at first. The extra expense of remelting matte 
will be about $2 per ton where coke is $12. There is a small 
loss by flue dust in addition, so that on 50 per cent, matte the 
remelting expense would be $i per ton of copper produced, or 
0.2 cent per pound of copper converted, which is approximately 
the difference of cost in favor of a plant where the matte is 
taken directly from the smelting furnaces to the converters. 

This cost would be much less in a locality where labor and 
fuel are cheaper than in Montana, but in most places in the 
West the figures would apply, and the cost per pound of copper 
would also increase rapidly as the grade of the matte fell 
below 50 per cent., or the tonnage decreased. For instance, 
when the tonnage handled at the converter plant was small 
the items of general expense were just as much as when it 
was large, and would greatly affect the result for that month. 

The average cost for converting 55 per cent, matte may be 
stated at 0.65 cent per pound of copper, divided as follows : 

Remelting matte 2 

Labor and lining for converters 25 

Labor on converters 1 

Resmelting converter slag 05 

Supplies 05 

.65 
For a plant without remelting 45 

The losses in converting as shown at Anaconda were about 
3 per cent, of the copper contents, but there was a wide 
difference between assays of samples of matte taken at the 
converter i)lcint and at the smelter. This difference amounted 
to 0.5 per cent, on all the matte received, and represented 
about 1 per cent, of the loss, so that the real loss was probably 
only 2 per cent, of the copper. 

The silver loss was apparently high according to assays of 
the converter copper, but the casting showed a corresponding 
gain, and after making allowance for this the loss of silver in 
converting was reckoned at less than 1 i)er cent. 



CHAPTER XIII. 
Blowing a Converter Charge. 

The operation of blowing a charge in a converter is one 
requiring much experience and attention on the part of the 
skimmer, and can only be learned by actual practice in the 
business. A knowledge of the changes of flame coloration at 
the nose, indicating the condition of the charge, is onh^ attained 
after the apprentice has given much attention to it, and in 
some cases where color-blindness may exist it is impossible 
for him to do so at all. The first portion of the blow is 
usually from 40 to 60 minutes' duration, although if the charge 
be too heavy it may be prolonged for much longer time. 
The flame is a light green color, with an occasional shade of 
yellow in the first stages, probably due to volatilized sulphur. 
As the time for skimming approaches occasional flashes of 
azure blue may be seen mingling with the light green, and if 
sufficiently prolonged it will become wholly azure blue. The 
converter must be turned down and the blast shut off before 
this change in the flame coloration has gone too far, or the 
entire contents of the converter will frequently be blown out 
and scattered over the building. The Avriter has seen charges 
weighing several tons foam and shoot matte thirty feet in the 
air from being overblown only a few minutes. 

The cause of these explosions is the oxidation of a portion 
of the copper which enters the slag, and when the vessel is 
turned or the equilibrium of the slag and matte disturbed, it 
results in the mixing of slag containing oxide with the white 
metal below ; the sulphur of the Avhite metal has a reducing 
effect on the copper oxide and the result is a sudden precipi- 
tation of copper and the formation of large volumes of SOo 
gas, which causes the charge to foam and at times throw the 

(74) 



BLOWING A CONVERTER CHARGE. 75 

greater portion of it out of the converter. This may happen 
before the converter is turned down owing to the agitation 
of the charge by the blast, but it is most active while the 
vessel is moving. It may happen after the charge has been 
skimmed, but in this case it is usually not so serious and 
shows that the skimming has been done too soon or before the 
iron has all been oxidized, that portion of the iron remaining 
in the matte having formed slag, which has absorbed copper 
oxide from the charge below and has been acted upon by 
the sulphur in the white metal. It is frequently necessary to 
skim a charge twice in order to remove the slag formed and 
take some of the burden off the blast. This will allow the 
blast to penetrate the charge faster and save time, as well as 
avoid the possibility of an}^ overblow. It also equalizes the 
charge burden in cases Avhere several converters are furnished 
blast from the same pipe. 

The formation of copper oxide will not take place to any 
great extent as long as iron is present in the matte, but as 
soon as the iron is all oxidized and gone into the slag as 
silicate, then copper begins to oxidize rapidl}^, and if the slag 
is not taken off before this begins it will unavoidably result 
in a large portion of the slag being thrown out. After the 
slag has been skimmed off the copper will be oxidized just 
the same, but so long as there is sulphur present in the 
charge the reduction goes on just as rapidly, so that there is a 
constant precipitation of copper and liberation of SO2. It 
occasionally happens with the best of skimmers that they will 
overblow a charge slightly and have slag shooting all over the 
building, but there is a sharp rivalry between them, and a 
rude sort of justice is meted out to the offender b}^ jeering 
and what is commonly called the " horse laugh." 

The writer has personally blown and skimmed many charges 
and has had just such experiences, and has learnt that many 
of the men who are able to handle a charge perfectly do not 
know the reason why certain things occur, but they are sure 
nevertheless of what will occur if nature's rules are violated. 

The slag should be poured off as quickly as possible and in 
a steady stream without moving the converter any more than 
is necessary, otherwise considerable matte may escape. The 
slag should flow easily and in a solid, dense stream, but if 



76 LEAD AND COPPER SMELTING. 

matte is escaping with it the bottom or back part of the stream 
will be seen to vibrate while the slag itself will be less fluid, 
flowing with the viscosity of molasses. At this point the rabble 
is shoved into the stream, and if the small crust of slag cut away 
shows matte the pouring is stopped and the remainder of the 
slag is skimmed off. Some slag will still remain behind, but the 
skimming should in all cases be as clean as is practicable under 
the circumstances. It sometimes happens, especially in large 
converters running on high-grade matte, that the slag will be 
granulated or only partially liquid. This may be due to a 
cold converter in case it is the first charge, or to the lack of 
heat developed by the charge due to slow running, which 
latter may be due to low blast or too large a charge. In such 
a case the remedy depends upon the cause. If it is due to a 
cold converter the addition of from 2000 to 5000 iDounds of 
matte will in most cases cause the slag to become liquid. If 
the charge is too heavy the blast pressure must be increased or 
a portion of the slag skimmed off in the best way possible, 
and the finish blow made without skimming clean. The un- 
fused material can be smelted by a small tap of matte after 
the copper is poured off, the slag poured and another tap 
taken to furnish metal enough to finish. With very low- 
grade matte several taps have to be taken to get the required 
amount of white metal to finish. At Aguas Calientes the 
converter was eight feet in diameter, the largest copper con- 
verter in the world, and on low-grade matte a charge of 8000 
pounds for the first tap would be taken and blown to white 
metal, the slag poured off, and a second tap of 10,000 pounds 
poured in and blown to slag and skimmed in the same way. 
A third, and sometimes a fourth, tap would be skimmed 
before the resulting white metal would be sufficient to stand 
above the tuyeres until the charge was finished. In this way 
charges of 45,000 pounds of matte were not uncommon, and 
the resulting copper would not be more than 40 bars of 200 
pounds apiece. With matte as low as 30 per cent, the con- 
verter increased in size so rapidly, owing to the corrosion of 
the lining, that it was difficult to get a charge to finish. After 
the vessel had finished the last charge that the lining would 
stand, a tap of from 10,000 to 15,000 pounds of matte would 
be put in and blown to slag, or until it was in danger of coming 



BLOWING A CONVERTER CHARGE. 77 

through the side or bottom. This was for the purpose of 
washing out the copper adhering to the lining after the last 
charge. The resulting white metal was dumped and used as 
scrap on subsequent charges. 

A considerable quantity of cold matte of high grade is 
necessary after the skim to keep the temperature of the 
charge from rising too high. If the charge is too hot the 
oxide of copper formed is not reduced before it forms a sili- 
cate with the lining, and as this is very infusible it is chilled 
by the blast on the tuyeres, resulting in clogging them and 
making the charge run slow and necessitating much punching. 
It seems paradoxical, but it is nevertheless true, that above 
the correct temperature the charge runs much slower and the 
tuyeres are hard to keep open. By the addition of white 
metal in large lumps the difficulty can be avoided. 

The use of low-grade matte to cool the charge results in the 
formation of more slag owing to the presence of iron, and this 
slag will, if the quantity is large, result in slow running and 
shooting out of the vessel as previously mentioned. If no 
scrap or white metal is to be had it is sometimes necessary to 
throw in a bar of copper to bring the temperature down. 
The question of temperature must of necessity be told by the 
appearance of the flame at the nose, and is quite as important 
as the other indications of slag and finished copper. 

In the first part of the blow, or before the slag has been 
skimmed off, it is not so much a question of temperature as 
how the converter is taking the blast. If the flame from the 
nose is the full size of the opening and appears to be leaving 
the converter with considerable velocity, it indicates that the 
tuyeres are open and the charge is working rapidly and satis- 
factorily ; but if it goes in puffs and has a choked and irreg- 
ular movement, the tuyeres need to be punched until they are 
free from obstruction. The tuyeres should always be iDunched 
before a charge is put in so that the blast may enter as freely 
as possible from the beginning. If the charge is sufficiently 
hot when it enters the converter it will need very little 
punching, and will start off with a very dense discharge of 
smoke and considerable volatilized sulphur. Shortly after- 
ward, probably fifteen minutes, it will have slowed down and 
will need punching for the next five or ten minutes, when 



78 LEAD AND COPPER BMELTING. 

o\Ying to the increased temperature it will run freely until 
time to skim. If the charge is cold when it enters the con- 
verter, punching will be necessary from the beginning and 
for a longer time thereafter than if it were hot. After skim- 
ming the scrap, white metal should be thrown in, and any 
cleanings that may be on hand should be introduced to the 
extent of about 10 per cent, of the weight of the charge if the 
slag is liquid and hot, but in less quantity if it is viscid, and the 
blast should be turned on as quickly as possible. Punching 
will be necessary if the flame indicates that the converter is 
not taking air properly. As long as the flame continues to be 
voluminous and leaves the nose freely all may be well, but if 
it assumes a bright brassy color and slows down the charge 
is probably becoming too hot, when more scrap should be 
added according to the requirements, and the tuyeres punched 
until the proper action is restored. If the flame assumes a 
light orange color that gradually turns into a darker shade, 
and then takes on a copper bronze color, the indications are 
that it is rapidly approaching a finish, when very close 
attention is necessary to avoid an overblow and an oxidized 
charge. 

Just the proper point to turn the vessel down is reached 
when the little particles of copper ejected from the converter 
give the appearance of very fine gauze or lines of a copper 
color, and when coming in contact with any obstacle they 
cease to adhere, as they will continue to do so long as they 
are of matte. The copper adhering to the punch-rod will also 
indicate very closely the time of finish. 

The vessel is then turned down and the granulated slag 
which will be present on top of the charge is shoved aside by 
means of the rabble until the surface of the liquid copper is 
exposed. If on skimming off a clean surface the charge shows 
a bright metallic mirror of copper, the charge is poured into 
moulds, but if the surface is covered by a skin of black 
sulphide of copper it is necessary to turn on the blast again 
for a short time. 

If considerable slag is present, or if the charge is slightly 
overblown, the copper will be covered with a layer of slag 
which will foam and bubble and require some time and con- 
siderable cold slag or cleanings from the floor to chill it 



BLOWING A CONVERTER CHARGE. 79 

around the nose before the copper can be brought to view. 
If there is sufficient copper in the converter to stand above 
the tuyeres it is possible to completely oxidize a charge, but 
although this sometimes does happen, it is through careless- 
ness or a mistake in judgment on the part of the skimmer. I 
have known it to happen when copper oxide on a charge 
already slightly overblown was mistaken for matte and the 
charge was kept working long after it should have been poured. 
Such things will happen, but seldom more than once to the 
same man, and it is not by any means a sign of incompetency, 
and having once occurred will improve the future service of 
the skimmer. 

If towards the latter portion of the finish blow the flame 
becomes very dark and red the charge is becoming cold, and 
the probabilities are that it will not finish without the addi- 
tion of more matte. If it does finish cold it will be difficult 
to pour and will leave much copper adhering to the lining. 
In such cases another small tap of matte is put in and blown 
to slag and skimmed, or at times when the matte is unusually 
low it is necessary to throw in a large quantity of cold scrap 
for the purpose of chilling the slag and making it impossible 
to skim. The reason for this is that with low-grade mattes 
the corrosion is so rapid and the addition to the copper con- 
tents of the charge so small, that instead of increasing the 
height of the charge above the tuyeres it is probable that it 
would be decreased if the slag were poured off. So that it 
becomes necessary to granulate the slag and force it to mix 
with the copper, raising the charge above the tuyeres in order 
to insure the blast penetrating it until the copper is finished. 

The methods that are used to develop heat in a charge 
that has run cold are, first, the addition of billets of wood ; 
heavy cordwood is preferable on account of its greater density 
and the fact that it will, by floating in the charge and by the 
gas generated, raise the charge level so that the blast will 
continue to penetrate the metal bath and develop heat instead 
of blowing over the top and freezing it. Second, the addition 
of lump coal, the effect of which is the same as cordwood. 
Third, the addition of a small amount of matte, finishing 
with granulated slag as described above. The reason for 
granulating the slag is that in this condition it will not cause 



80 LEAD AND COPPER SMELTING. 

the shooting and foaming of the charge described when over- 
blown. This last method is, as before stated, employed in 
case the matte is of low grade, but if it is high grade, say 55 
per cent., the slag may be safely skimmed off, the resulting 
white metal being enough to compensate for the extra corro- 
sion of the lining and also to make up the shortage previously 
existing. If the converter is of the cylinder or Leghorn type, 
it can be turned back until the tuyeres are brought to a 
lower level and the blast forced to penetrate the charge, but 
with the great majority of converters this is not possible, since 
turning the converter back beyond a fixed point is p>revented 
by the smokestack into which the fumes are discharged. 
Only the experience and judgment of the skimmer is to be 
relied on in such cases, and the remedy must be applied which 
is best suited to the conditions. 

The relative heat-developing power of sulphur and iron is 
very strikingly illustrated by the action of different grades of 
matte. A charge of low-grade matte will smelt fully half its 
weight of granulated slag left in the converter, while a cliarge 
of high-grade matte will only add to the difficulty. 

The low-grade matte containing more iron has greater heat- 
developing power, as well as more basic action on the silicious 
slag, and will bring it to a jjerfectly fluid condition, while the 
high-grade matte containing less iron and making a more 
silicious slag will be unable to smelt the accumulation in the 
converter. The heat developed in the first portion of the 
blow, as Avell as the fact that the iron is all converted to oxide 
much sooner than the sulphur, shows that iron has a stronger 
affinity for oxygen than sulphur, and in combining with 
oxygen developes more heat than the latter. This is also 
proved by the action of the charge in becoming rapidly hotter 
in the blow from matte to slag and gradually colder after tlie 
iron has been oxidized and the slag skimmed off. It is a 
common error often repeated that the heat developed in 
copper converting is due entirely to the burning of the sul- 
phur, while the fact is that the more heat is developed by the 
burning of the iron. 

For this very reason it becomes difficult to convert mattes 
as high as 65 per cent. Cu on account of the decreased amount 
of iron and insufficient heat development. It i« apDarent at 



BLOWING A CONVERTER CHARGE. 81 

once that the presence of the iron is necessary, and that, 
being present, silica must be provided in the lining for it to 
act upon to form a fluid slag, and, further, that if the corro- 
sion of the lining is stopped the process will be defeated. The 
silicious lining is as much a part of the process of copper con- 
verting as magnesia lining is in the basic Bessemer treatment 
of phosphoric iron, and it is suicidal to attempt any other 
kind of lining, either water- jacketed or basic. The improve- 
ment, if there is to be any, is to be in the line of mechanical 
devices, and the use of silicious ores to replace the expensive 
quartz and clay linings. 

At Aguas Calientes the lining was made entirely of ore, 
and this contributed a great deal to the success of converting- 
low-grade matte at that point. If quartz had been used 
instead of ore the expense would have been too great to admit 
of 80 per cent. Cu matte being converted. It was very 
fortunate that an ore of such ideal composition for the 
purpose was to be had. This ore came from Pachuca, State of 
Hidalgo, and was mined there in large quantities. 

A partial analysis showed: Si02 72 per cent., FeO 5 per 
cent., CaO 0.6 per cent., AI2O3 15 per cent. The ore ground 
in a Chilian mill with water was very plastic and did not 
need the addition of clay, and it was possible to run with a 
lining on which there was a margin of $20 (Mexican) per ton. 
If such ores could be obtained in Montana it would jDrove a 
bonanza to the copper converters. 

There are three kinds of finish on converter copper, accord- 
ing to the time the charge is turned down. 

The first shows a small amount of regal and is usually about 
95 per cent. Cu, and sometimes expands to such an extent on 
cooling as to make it exceedingly difficult to get the bars out 
of the moulds. On this account, as well as the extra time 
and expense in refining before casting into anodes, it is 
seldom made, and then only when a charge is too cold to be 
kept longer in the converter, or on account of weak lining. 
The second and most common and desirable of the three is 
called gas finish, on account of the large quantity of SO2 which 
leaves the metal on cooling. This finish shows no regal but 
contains SO2, dissolved in the copper to such an extent that a 
mould filled with the metal will, after the ebullition has 



82 LEAD AND COPPER SMELTING. 

ceased, not be more than one-third to one-half full, and it is 
necessary to pour into the moulds two or three times in order 
to make a fair-sized bar. 

The SO2 will remain with the copper as long as it is in a 
molten condition in the converter, but as soon as it strikes the 
cold mould and begins to solidify, the gas comes off rapidly, 
and if great care is not taken in pouring into the moulds, the 
copper will effervesce and run over like soda water. 

At times a crust will form over the top of the bar before 
the gas has escaped from the liquid interior, and then a 
rupture will take place and a stream of molten copper may be 
thrown a distance of several feet by the escaping gas. Serious 
burns frequently and unfortunately occur on this account, 
and it is rather dangerous to stand near the moulds until the 
copper is thoroughly solidified. There seems to be a very 
strong resemblance between the affinity of molten silver for 
oxygen and this peculiar action of converted copper and SO'2. 
It does not occur with blister made in reverberatories for the 
reason that all the oxidization takes place on the surface, 
while in the converter it goes on much faster and all through 
the metal bath, some of the SO2 formed being dissolved in the 
copper. 

The third is called blister finish and exhibits the charac- 
teristic blisters on the surface of the bars from which it gets 
its name. In a charge of ordinary size there is about ten 
minutes' difference in time between the first and second finish, 
and about five minutes between gas finish and blister finish. 
The blister finish contains a small amount of gas, but usually 
not enough to cause the copper to decrease much in volume 
on cooling. It is seldom that a whole charge, especially if it 
be a large one, will be blister finish. Usually the last few 
bars are gas finish, while the copper first poured from the top 
of the charge will be blister. In order to produce blister it is 
necessary to overblow slightly, and some copper oxide will be 
floating on the surface as slag. 

The distribution of silver in the copper bars, as determined 
by assaying samples taken from different i^arts, shows the 
folly of trying to get a correct sample excei3t after remelting 
or from the stream as it comes from the converters. The 
results of two bars sampled at Aguas Calientes are given below : 



BLOWING A CONVERTER CHARGE. 



83 



Blister Finish. 


Gas Finish. 


No. 


Part of bar. 


Assay in troy 

ounces silver 

per ton. 


No. 


Part of bar. 


Assay in troy 

ounces silver 

per ton. 


1 
2 
3 
4 
5 
6 
7 


End, 
End, 
Top, 
Bottom, 

Side, 
Side, 
Fin, 


358.1 

403.4 

608. 

366.2 

393.9 

372.5 

423.6 


1 

2 
3 

4 
5 
6 

7 


End, 

End, 

Top. 

Bottom, 

Side, 

Side, 

Fin, 


231.2 
247.4 
451.7 
203.6 
252.9 
235.9 
351.4 



In order to get a sample of a carload lot of converter copper 
it is necessary to take a sample of each charge and mark the 
number of bars on the ticket that is to go with it. The weight 
of each charge separately would be the correct way, but as 
the bars will average about the same, this can be used instead. 
A number of these charges are bunched together to make a 
carload, and the samples are cut up into small pieces and as 
many grammes taken from each as there were bars in the 
charge. These weighed portions are put together in a large 
clay or graphite crucible, melted in a blacksmith's forge, and 
granulated by pouring slowly on to a board set obliquely in a 
bucket of water. The stream of copper should strike the 
board about six inches above the water, and should fall a dis- 
tance of about two feet before striking. It will glance off in 
fine shots and, chilled by the water, will be found as bright 
granules when the water is poured off. 



CHAPTER XIY. 

Design of Converter Plants. 

When copper converting was first started in this country 
the converters were made stationary, and in order to reline 
them they were filled with water, as described for the old 
plant at Anaconda. But the experience gained showed that 
this practice would have to be abandoned. As it was im- 
practicable to wait for them to cool in the stands, it became 
necessary to remove them and place a fresh vessel in place so 
that work could go forward without serious interruption. 

The Parrot plant was constructed with converters 60 inches 
diameter by 8 feet 6 inches high, which were removed on a. 
car. The lifting device was four jackscrews, one at each 
corner of the car, and the heads of these screws impinged on 
lugs riveted to the converter. The converter was turned on 
its back by the hydraulic cylinder, the car run under the con- 
verter, and the jackscrews applied until the weight of the 
converter was lifted off the trunnions. The trunnions were 
then uncoupled at the flange joint, one on each side between the 
vessel and the stand, the tips of the bearings remaining in the 
stand. The car bearing the converter was then run out to the 
relining house and another car with a converter lined and 
dried was run in place, and after coupling the trunnion tips was 
ready to receive a charge. This method was discussed for the 
Anaconda plant, but it was fortunately decided that a travel- 
ing crane would be better, and the plant was built according 
to the plans reproduced in the appendix. 

One of the difficulties encountered in this plan was to get 
the converter and cupolas to deliver their smoke into the same 

(84) 



ANCHOR FOR CONVERTER 
LINING, TO BE RIVETED TO 
•- INSIDE OF SHELL, ABOVE 
LINE A-B AT DISTANCES OF 
10"lN ALL DIRECTIONS FROM 
CENTER TO CENTER. TO 
> BE MADE OF STEEL. 




15 TUYERE HOLES 
REQUIRED 



Fig. 20.— Copper converters at Aguas Calientes. 



86 LEAD AND COPPER SMELTING. 

flue. It had previously been the practice to have the conver- 
ters turn down towards the cupola to receive the charge and 
turn up and blow away from it. This would have been im- 
possible with a traveling crane, and the difficulty was over- 
come by putting two converters to each cupola and making 
them turn down, away from the cupola, to receive the charge 
and blow towards it into the same flue. 

The first section of the spout from the cupola well was 
straight and emptied into a broad section which was movable 
and divided into two spouts, one running to the right and 
another to the left-hand converter. These spouts were 
mounted on wheels and could be swung around to either con- 
verter when a charge was needed. By means of a few shovels 
of clay the matte could be diverted to either spout, which 
would carry it to the converter. The trunnion coupling 
which was selected for use after many designs had been made 
is worth special consideration, since it is very strong in con- 
struction and can be coupled up or uncoupled in about 30 
seconds. The trunnion ring remains on the converter and is 
taken out with it. The tips of the trunnion ling are conical 
in shape, with the large ends farthest from the vessel. Thesfe 
cones fit into pockets in the trunnion tips, and are fastened to 
them by heavy gibs and keys, as well as bolts on the bottom 
(see Plate Y, appendix). 

The dimensions of the vessels adopted were 72 inches 
diameter by 10 feet high. It would probably have been 
better had they been made 6 feet 6 inches by 32 feet high, 
but at that time the large converters at Great Falls were not 
giving satisfaction, although later they improved wonderfully 
under diflferent management. The discouraging reports cir- 
culated had the effect of placing a limit on the size of the 
vessels at Anaconda. 

Since that time it has been the experience of the writer 
that very large converters do not do as good work on high- 
grade matte as smaller ones, while on low-grade mattes they 
do much better. The extremes of size thus far in use are : 
Parrot, 58 inches diameter by 8 feet 6 inches high ; Aguas 
Calientes, 96 inches diameter Iby 16 feet high. The converters 
at Aguas Calientes are illustrated in detail in Figs. 20 and 21. 
The general arrangement is shown in Figs. 22« and 22 J. 





IRED. 
ETEDTO 
SHELL TO CORRESPOND 
IN SIZE, ALSO BOLT 
HOLES TO CORRESPOND. 



anfi^^^^^rS^ 






CORE FOR 1 RIVETS 







BOLTS TO HOLD 
CLIPS TIGHT 



16 REQUIRED OF 

CAST STEEL 
CORE FOR IJ^'bOLTS 



i\° I T ffarfe 



TO FIT SHELL 



-12- 



THESE ARE TO COVER TUYERE OPENINGS 
AT "O" ON OUTER SHELL; 15 REQUIRED 
WITH BOLTS, ALSO HOLES IN SHELL TO 
BE THREADED TO FIT. 




•-A- 


r*r-i 












1 — ' 












^ 








"^ 







o 




o -* 


-2 


< 


.^_ 


o 


_ 





T 


< 12--H 












M 




PATTERNS FOR THESE 
BRACKETS MUST BE 
FITTED TO CONVERTER 

SHELL. 

4 REQUIRED 



FIG. 21. 

8'o"dia. copper converter 

CAPACITY 5-7 TONS. 

LA GRAN FUNDICION 

CENTRAL MEXICANA. 
VVGUAS CALIENTES, MEXICO. 



88 



LEAD AND COPPER SMELTING. 



The main objection to large converters on high-grade matte 
is that the slag is very frequently granulated and cannot be 
poured off, while with the smaller converters, as at Anaconda, 
no such difficulty was ever experienced ; in fact, the trouble 
was to get the matte high enough in grade so that the linings 
would last six or seven charges. 




.s'xVl^s'P- 



SECTION A-B 
Fia. 22a.— GENEEAIi ARRANGEMENT OF 8-FOOT DIAMETER COPPER CONVERTER. 

As the size of the vessel increases, the thickness of the 
lining also increases rapidly, and since it is very porous the loss 
of blast through the lining also increases. A point would 
soon be reached where with increased thickness of lining 
the decreased efficiency of the blast would more than counter- 
balance all the benefits of large vessels. The loss of blast 
through the lining depends a great deal on how the latter is 
put in. If it is tamped in around a cone the loss will be less 
than if it is simply built up by hand. But tamping linings 
around a cone is much slower and more expensive, and tests 
made on the life of such linings did not show that they lasted 
any longer than those put in by simply pounding the wet 
material into place with the hands. 

There are two methods of arranging a converter plant, 



DESIGN OF CONVERTER PLANTS. 



89 



either of which is good, and the choice between them depends 
upon local conditions. If the tonnage is large and a great 
quantity is to be turned out, the crane plant is by far the 
better, and the size of the vessels should not be less than 7 
feet in diameter by 13 feet high. If the tonnage is small the 
plant should be constructed on about the same lines as the 



Hi ■ 46 left 



H Plate \\ 



I5''(3dt, 




Fig. 226.— GENERAii arrangement of 8-foot diameter copper 



CONVERTER. 



90 LEAD AND COPPEK SMELTING. 

Aguas Calientes plant, but with the joint in the lining at the 
top of the trunnion ring instead of at the bottom. Of the 
crane plant nothing in addition to what has been said of 
the Anaconda is necessary, but of the handling by cars at 
Aguas Calientes this can be said : that the plant will cost much 
less to construct and a much larger converter can be used 
than could be handled by any crane that a converter plant 
could afford to erect. The vessels at this plant, when freshl}^ 
lined, would weigh fully 40 tons, and the buildings and crane 
necessary to handle such a great weight w^ould cost fully 
$30,000 more than the plan adopted. 

The practice at this plant was similar to that of most steel 
plants where the vessels and bottoms are removed on a car, 
the car with the vessel on it being lifted into position by a 
hydraulic piston about 15 inches in diameter, acted upon by 
water pressure of about 500 to 600 pounds to the square inch. 
The top end of the piston is fitted into a very strong frame, 
which carries a section of the track long enough to allow the 
wheel-base of the cars to be moved forward or backward a 
few inches without running off the end of the rail. The con- 
verter is put into place in two sections. The top section being 
cleaned of slag about the nose, is then taken by the car on to 
the elevator below the trunnion ring, hoisted into place and 
bolted to the ring in an inverted position. It is then turned 
upright by the hydraulic cylinder attached to the trunnion 
ring and the bottom section fully lined and dried, run under 
and hoisted into position and bolted to the trunnion ring as 
well as to the top section. The remainder of the lining must 
then be put into the nose, and as this is a very long, tedious 
job the less of it there is to do the sooner it will be done. 
The mistake was made of cutting the converter in two below 
the trunnion ring instead of above or at the upper edge. The 
bottom section should be the largest piece of the shell, so that 
as much lining as possible may be put in before the top is put 
on. The idea in constructing the plant in this way was that 
the bottoms were the only iDortion of the lining that were cor- 
roded, and that they could be removed and renewed just as is 
done in the steel business. This idea, of course, is wrong. The 
lining is corroded wherever the slag and matte touch it, ex- 
cept right at the nose, where the slag tluit is thrown out freezes 



DESIGN OF COPPER CONVERTERS. 



91 



and causes it to grow smaller. Owing to the unfortunate cir- 
cumstance that a lining 2^ feet thick, of clay and quartz, with 
nothing to support it, will fall out, the top section could not 
be lined bottom side up and then turned upright to receive 
the bottom without losing all the lining in the upper section 



SECTION AT A-B 

21" 

18 



SECTION AT C-D 
13M- 




FiG. : 



SECTIONS OF COPPER MOULDS. 

-Cope and core apparatus for making copper moulds. 



and causing a wreck that would take several hours to repair. 
If, as stated, the joint had been made at the upper side of the 
trunnion ring instead of at the lower, the lower section would 
have been increased by that amount and the quantity of 
lining necessary to be put in after the cap had been put on 
greatly reduced. Owing to the length of time necessary to 



92 LEAD AND COPPER SMELTING. 

pass 15,000 pounds of lining in at the nose and iDut it into 
place, as well as the time required to dry the lining and heat 
the vessel to a point where it will be in a fit condition to 
receive a charge, the time in which a change of vessels could 
be made was unusually long. From ten to fifteen hours were 
required as against five minutes at Anaconda from the time 
a charge was poured until another could be run in. If the 
joint had been in the proper place a change could have been 
made in about four hours, including firing the fresh lining. 

In other respects the plant was a good one, and the large 
vessel was certainly a great improvement over a smaller one 
for the treatment of low-grade mattes. It also made it pos- 
sible to use for lining material silicious ores, on which there 
was a margin of profit, which would have been totally unfit 
for use in smaller vessels. 

The use of cast iron or steel moulds for converted copper is 
rather exijensive, and it has been the practice at many places 
to make the moulds of copper. This is done by pouring 
refined copper into a mould made of two ells clamped together, 
and then plunging a core-bar into the metal bath, which 
would make the mould. It involves considerable expense for 
labor and supplies, and the experiment was tried of making 
the moulds direct from the converter. The moulds could be 
made very easily by means of the cope and core apparatus 
shown in Fig. 23, but when in use the stream of copper from 
the converter would strike on one part of the bottom until it 
became hot enough to weld, and the result w^ould be that the 
bar and mould could not be separated on account of a union 
covering a space of about 1^ inches in diameter. If the 
copper being poured was very hot it would bore holes into the 
copper moulds as quickly as hot water poured upon ice, and 
even after the difficulty with the bottoms of the moulds had 
been remedied by means of the cast-iron plate placed on top 
of the core before casting, great care had to be taken that the 
stream did not impinge on the sides of the moulds. 

The cope and core were placed on top of an ordinary slag 
pot, preferably one the bowl of which was cracked, so that it 
could remain there without being changed. The cast plate 
shown was then placed on top of the core and the pot run under 
the converter nose. The copper Avas poured in slowly to allow 



1 



DESIGN OF COPPER CONVERTERS. 93 

the SO2 gas to escape and the space between the cope and 
core to fill up solidly. The pot was then pulled to one side 
and the cap lifted off by means of the traveling crane, and a 
bar with chisel point driven between the copper mould and 
the cast core at one end. A couple of men would then pry 
down on the bar until the mould would come away from the 
core, or until the mould and core were raised from one side 
of the iDot, when a few sharp blows from a sledge on the raised 
end of the core would bring them apart. The cast iron would 
come away with the copper mould and would serve the double 
purpose of protecting the core while making the moulds and 
preventing the welding of bars to the moulds while in use. 

This scheme worked very satisfactorily, and copper moulds 
Avith cast-iron bottoms were used entirely. It was necessary 
to make from 12 to 20 each day owing to their breaking after 
about five to seven days' use. The scrap or broken ones were, 
of course, available as bars for the casting furnaces. 

The life of these moulds depended a great deal on the kind 
of finish copper that was used in making them, moulds made 
from blister finish being found much better than gas finish. 
While making moulds two or three cores on pots would be 
used at one time, one filling while the others would be strip- 
ping or being washed in limewater to prevent them being 
burned. By the use of a few extra men and a traveling 
crane twenty moulds could be made in an hour. 



CHAPTER XY. 

Lining a Converter. 

It IS unfortunate that in most places the quartz and clay 
are so entirely distinct. If, as at Aguas Calientes, they are 
both contained in one ore the lining is much more homogene- 
ous and compact. The linings in Montana are made from 
quartz which has no plastic or adhesive qualities and is very 
refractory, and a fat, sticky clay, which is not refractory and 
which melts away from between the particles of quartz, allow- 
ing them to drop off and mix mechanically with the slag or 
copper. This probably accounts for much of the difference in 
the composition of the converter slag at Anaconda and Aguas 
Calientes, which averaged as follows : 





SiO^. 


FeO. 


Cu. 


Aguas Calientes 

Anaconda 


25 per cent. 
36 percent. 


62 per cent. 
49 per cent. 


5 per cent. 
5 percent. 



The matte at Anaconda would average 55 per cent, and at 
Aguas Calientes not more than 35, and part of the difference 
in slag composition is due to the fact that high-grade matte 
makes more silicious slags than low grade. 

The introduction of silica through the tuyeres was not 
attempted at Anaconda, for the reason that it was apparent 
to the writer that so much silica as would be required could 
not be blown in within the short time allowed, and, second, 
because it would act as a sand blast and ruinously cut into 
the ironwork of the converter in a very short time. 

Even if the silica could be introduced in this way it would 
still produce another trouble in the converter. The small 
particles of silica being cold would combine very imperfectly 
with the iron oxide and produce granulated and pasty slags, 
which could not be skimmed or poured out of the vessel. 

(94) 



LINING A CONVERTER. 95 

About 60 tons of quartz and 7 tons of clay were consumed 
daily for lining, and two sets of crushers and rolls and three 
Chilian mills were kept going day and night to grind the 
material for lining. The mixture used in charging the Chilian 
mills was 40 shovels of quartz to five of clay. This was 
ground together with hot water in winter time, and cold when 
the weather would permit, until the whole mass was reduced 
to the consistency of a stiff mud and the quartz pulverized to 
the size of peas. In this condition it was discharged from the 
mills into a stock-pile or direct into wheelbarrows to be taken 
10 the converters in process of lining. Before passing to the 
liner it was pounded with the shovel until it adhered thor- 
oughly together, and was then cut into cubes by the shovel and 
passed into the converter. As soon as the converters were 
removed from the stands the nose section was taken off by 
removing the keys from the key bolts and lifting with the 
crane, when the joint Avould generally break, but if it did not, 
a wedge would be driven in and a fracture started. To hasten 
the cooling the interior would be sprinkled with water from a 
hose. The skin of slag was then cut out and the bottom sec- 
tion relined, after which the top was put on by the crane and 
the lining finished as high above the joint as possible. The 
converter was then taken by the crane to a place where there 
was a blast connection through a three-inch hose to a branch 
from the main supplying the blast furnace. A fire was 
kindled with oily waste and dry wood, and shortly afterwards 
coke was put in and the blast turned on. The fire was kept 
up until the converter was required for use in the stands. 
From three to four hours was required to dry and bake the 
lining sufficiently to insure it from falling out when the vessel 
was inverted. It was found that if the fire was allowed to go 
out the lining would contract so much that the probabilities 
were it would collapse when put in the stand and turned 
upside down. 

The life of a lining should not be measured so much by the 
time or charges as by the copper produced and the matte con- 
verted. The production of a lining is dependent on the com- 
position of the clay and quartz, as well as upon the grade of 
matte converted. If the clay is not plastic or the mixture 
poorly ground, the lining may collapse after a single charge. 



96 LEAD AND COPPER SMELTING. 

On the other hand, if all things are working to the best 
advantage and the grade of the matte will average 55 per 
cent. Cu, the first charge for a six-foot vessel should produce 
about eight to twelve bars of copper of about 250 pounds to 
the bar ; the second charge from ten to sixteen bars, the third 
from sixteen to twenty, and the fourth from eighteen to 
twenty-eight, the fifth from twenty to thirty, the sixth from 
twenty to thirty-five, and so on until the lining becomes too 
thin to stand another charge. A production of 100 bars is a 
good run for a lining on 55 per cent, matte, and all statements 
to the contrary must be taken with considerable doubt. In 
one work on the subject a statement is made that the linings 
are usually exhausted after the ninth charge. It is safe to 
say that with the size of vessels in use at the Parrot works, 
where this remarkable work was done, all the linings that have 
made nine charges in the past two years can be counted on 
the fingers of one hand. At Aguas Calientes, where the 
vessel was 8 feet in diameter and the lining 2J feet thick at 
the tuyeres, one charge of 30 per cent, matte weighing 40,000 
pounds would, if added 8,000 to 15,000 pounds at a time, 
finish about forty to forty-five bars and corrode the lining so 
much that a second charge could not be finished. With such 
low-grade matte the second charge would only be blown to 
white metal and skimmed, the 80 per cent, copper matte 
being poured into beds and returned to the next charge as 
scrap. A single lining would, including the washout, convert 
about 50 tons of 50 per cent, matte. 

The experiment was tried at Anaconda of turning the con- 
verter on its back in the stand and putting in a large patch 
of green lining wherever needed, and allowing it to dry 
thoroughly before a charge was run in. By repeating this 
operation several times twelve charges were finished with a 
total production of 212 bars for a converter 60 by 60 inches 
square. Aside from the large vessels at Aguas Calientes this 
is the largest production for a single lining that I know of. 
The largest charges thus far finished were probably made at 
Great Falls, where according to report something over 80 bars, 
or about 16,000 pounds, of copper have been poured. The 
largest single charge at Aguas Calientes was 75 bars. 



CHAPTER XYI. 

Casting Anodes Direct from Converter. 

The casting of anodes from the converter had been attempted 
at the old plant at Anaconda, and was successful enough to 
indicate that it could be done by the assistance of a traveling 
crane. When the new plant was well started and the men had 
become accustomed to the use of new appliances, the casting 
of anodes was carried on for some time, although eventually 
abandoned because of the objection made that the impurities 
in the converter anodes caused the electrolyte of the refinery 
to become too impure, it being stated that the cathode copper 
from such anodes would be low in conductivity and unfit for 
wire bars. However, it was demonstrated that anodes of 
reasonably uniform weight and density could be cast at a 
saving of about 0.35 cent per pound, or $7 to the ton. The 
anodes were cast on edge between cast-iron moulds grouped 
together so that the face of one mould w^ould be the back of 
the next, the set being held together with iron clamps with 
springs to allow for expansion when filled with copper. Owing 
to the rapid chilling of the copper when it was poured into 
the moulds, it was necessary to be able to pour along the 
entire length of the mouth of the mould, as well as to change 
quickly from one mould to the other and back again. The 
escaping SO2 gas would cause the copper to shrink in the 
moulds, and the anodes would be hollow unless filled a second 
time. To overcome these difiiculties a car with a movable 
table-top on diiferential rollers was designed, and the set of 
moulds put on this table. All the peculiar difiiculties were 
overcome except the warping of the cast-iron moulds by the 
heat from the copper. This caused the anodes to be somewhat 
irregular in thickness and weight, owing to the fact that the 
moulds after being used for a few days would buckle so much 

(97) 



I 



98 LEAD AND COPPER SMELTING. 

that they could do longer be drawn close together by the 
clamp. This variation would amount to 20 per cent, of the 
weight of the anodes, in moulds made to cast plates of equal 
thickness. The anodes were taken from the moulds to a large 
Gate shear, where the ragged upper edge was sheared off, and 
two holes punched in the ears by which they were suspended 
in the tanks. 

It was found that plates could not be cast less than 1 inch 
thick on account of the chilling of the copper before the 
mould had been filled in all parts. Anodes 1^ inches in 
thickness of the dimensions required would weigh 230 pounds 
made of converter copper, and if made of cast copper would 
weigh considerably over 300 pounds, the difference in weight 
being due to the porous character of the converter copper 
caused by the escaping SO2 gas. Anodes of unpoled and con- 
verter copper, which contain more impurities than were con- 
tained in these, are used at other refineries, but with what 
success as regards the conductivity of the cathode copper I am 
unable to state. 

The average assay of samples of converter copper showed 
99 per cent. Cu, with silver varying from 80 to 120 ounces and 
gold from two-tenths to five-tenths ounce to the ton. There 
was considerable SO2 gas retained in the pores of the anodes, 
which it was stated was converted into sulphuric acid in the 
electrolyte, resulting in the increase of acidity of the solutions 
and rendering unnecessary the addition of free acid. This, as 
well as the porous character of the converter anodes which 
later would present greater surface to the action of solutions, 
should be considered as points in their favor. 

If the anodes are cast from the converter in open moulds, 
as at Great Falls, the thickness and weight are subject to 
greater variation than if cast on edge. The stream of copper, 
as it comes from the converter, has to be broken up and 
deflected to a different part of the mould by allowing it to fall 
on a board held by an attendant. If this is not done the 
copper will set before the lugs have been made, and the 
anode will be much thicker in the middle than on the edges. 

There is a much greater production of scrap in the refinery 
from converter anodes than from cast anodes, OAving to the 
imperfections of the former. 



CHAPTER XYII. 

Cost of PROcucr^G Copper at A^^AcoxDA. 

In exxDlanation of the following figures, in case they should 
not seem clear and intelligible, it is only necessary to state 
that the losses in percentage are on a basis of the material 
charged to the different departments. To get the copper 
marketed it is necessary to deduct them in their order from 
100 per cent, in the ore after multiplying by the percentage 
delivered to that department. 

For example, if 18 per cent, is lost in dressing and 9 per 
cent, in smelting, then 100 — 18 = 82 per cent, delivered to 
smelter; 82 X 9 per cent. = 7.38 per cent, of Cu in the ore 
lost in smelting; 82 per cent. — 7.4=74.6 delivered to con- 
verter ; 74.6 X 3 = 2.238 per cent, of Cu in ore lost in con- 
verting ; 74. 6 — 2.2 = 72.4 delivered to casting ; 72.4 X 1 = 0.72, 
and 72.4 — .7 = 71.7 delivered to refinery, etc. 

In the same way the costs are figured, only the cost for the 
department is divided by the per cent, of Cu marketed. Start- 
ing at 100 per cent, and working backwards, 1 per cent, loss 
in melting would have the cost divided by 99 per cent., 1 per 
cent, loss in casting by 98, etc. 

Copper in ore 100 per cent. 

Loss in dressing 18 

Delivered to smelter 82 

Loss in smelting 9 

82 X .09 = 7.38 ( say 7.4) ; 82 — 7.4 = 74.6 

Copper delivered to converter 74.6 

Loss in converting 3.0 

74.6 X .03 = 2.238 (say 2.2) ; 74.6 — 2.2 = 72.4 

Copper delivered to casting department 72.4 

Loss in casting 2.0 

72.4 X .01 =0.724 (say 0.7); 72.4 — 0.7=71.7 

Copper delivered to refinery 71.7 

Loss in refining 0.5 

72.4 X .005 = 0.3585 (say 0.3) ; 71.7 — 0.3 = 71.4 
(99) 



100 LEAD AND COPPER SMELTING. 

Copper delivered to melting department 71.4 

Loss in melting 1.0 

71.4 X 0.01 = 0.714 ( say 0.7) : 71.4 — 0.7 = 70.7 

Copper finally recovered from ore 70.7 

Cost of dressing 0.53 per pound of copper in concentrates. 

Making the calculation in the same manner on the basis of 
the copper contents of the concentrates delivered from the 
dressing works, it appears that 86.2 per cent, is recovered, L e., 
the loss in smelting and converting is 13.8 per cent. The cost 
of dressing per pound of copper marketed is consequently : 
0.53-1-0.862 = 0.614. In a similar manner the cost of smelt- 
ing per pound of copper marketed works out : 2.035-7-0.945 = 
2.153; cost of converting matte to blister: 0.6870 -r- 97.5 = 
0.705 ; cost of casting : 0.35 -^ 98.5 = 0.356 ; cost of refining, 
1.00 cent; cost of melting, 0.40 cent, cost of mining: 2.2 -r- 
70.7 = 3.112. The recapitulation is as follows : 

Cost mining per pound Cu sold 3.112 cts. 

Cost concentrating per pound C a solJ. . . 614 

Cost smelting per pound Cu soL- 2.153 

Cost converting per pound Cu sold 705 

6.584 , 

Cost casting per pound Cu sold 356 

Cost refining per pound Cu sold 1.000 

Cost melting per pound Cu sold 400 

Total cost per pound Cu sold 8.340 

To each pound of copper there is recovered an average value 
of about 4 cents in precious metals. 

The building at Anaconda was designed for twelve con- 
verters running and thirty-six shells, being three shells to 
each stand. There were to be six cupolas, one to every two 
stands, and two blast furnaces to work over the converter slag. 
Three No. 7 Roots blowers were required to furnish the blast 
for the cupolas, blast furnaces, and drying out the converters. 

There were four blowing engines with a capacity of 2,500 
cubic feet each per minute and two with a capacity of 8000 
cubic feet each, making a total capacity of 26,000 cubic feet 
per minute, or about 2,200 cubic feet per minute for each 
converter in operation. 

To furnish the steam there were eight firebox boilers with 
shells 72 inches by 18 feet besides the firebox. Each had a 
steaming capacity of 160 indicated horse-power, making 1,280 



COST OF PRODUCIXG COPPER AT AXACOXDA. 101 

horse-power in order to turn out 11,000.000 pounds of copper 
per month. This would be the maximum if all were running 
at one time, but out of twelve converters seldom more than 
eight were running, and as the speed of the engines was 
regulated automatically by the air pressure, the maximum 
power was seldom required. The blowing engines were all 
furnished with Corliss steam valves, and the four small ones 
with Corliss air valves. The large engines were built accord- 
ing to specifications with gridiron slide valves for the air 
cylinders, and it was found that these were much better than 
the Corliss valves. The cost of this plant was about $400,000. 



APPENDIX. 

SPECIFICATIONS OF BUILDINGS AND MACHINERY FOR COPPER CON- 
VERTER PLANT, ANACONDA MINING COMPANY, 



ANACONDA, MONTANA. 



All the machinery described in the following specifications 
is to be first-class in every respect, both as regards material 
and workmanship. The machinery described is to be com- 
plete as per specifications and blue prints furnished, omissions 
in specifications notwithstanding. 

Buildings : General appearance and dimensions as per 
Plates I and II. 

Note : Dimensions given are from G to G of posts. 

Stockhouse : 324 ft. long, 23 ft. wide, and sujDported on 
posts 12 ft. G; on one side these posts are 18 ft. high ; on the 
other side they are 35 ft. high. The 18-ft. posts are attached 
to and stand on 24-in. deep plate girders ; these girders are 
securely fastened to the 85-ft. posts, and are to be i^rovided 
with angle irons or knees riveted to top of girders a suitable 
distance apart to fasten 12 x 14-in. stringers to for a standard 
track (4 ft, 8 in. gauge) ; track to be in center of building. 
Roof to have a longitudinal opening 4 ft. wide the whole 
length ; the upper side to be provided with window openings 
— two windows, twelve 10 x 18-in. lights in each bent. The 
lower side towards the Converter Building to be left open. 

Cupola and Converter Building : 240 x 70 x 26 in. Trusses 
are attached to the 35-ft. posts of the above-mentioned build- 
ing in the rear, and in front on 26-ft. high posts 12 ft. G. 
These posts, or a suitable number of them, are provided 
with brackets to carry a girder for traveling cranes. Roof to 
be covered, except the nine ventilator holes 12x17 ft. 6 in., 
and the holes for six cupola stacks 12 x 8 ft. ; a frame to be 

(102) 



SPECIFICATIONS. 103 

made around respective openings and their sides to be covered, 
and a suitable flashing made between sides and roof. The 
ends of this building to be covered on a suitable framework 
to within 8 ft. from the ground, except where adjoined by 
buildings. 

Blast Furnace Building : 84 x 36 x 29 ft. high. Trusses are 
attached to posts of Stockhouse Building in the rear and on 
29-ft. posts 12 ft. C in the front. Roof to have three openings 
8 X 12 ft. for furnace stacks, located as per drawings. Roof 
and sides to be covered within 8 ft. of the ground. 

Silica Mill Building : 72 x 56 x 20 ft. Trusses are supported 
on one side on a stone wall ; wall plates and anchor bolts for 
this. On the other side they are supported on 20-ft. high 
posts 14 ft. C. One end and one side are provided with 
window openings, as per drawing. Roof has two dormers - 
14 ft. wide 8 ft. high, located as per plan. Roof and sides to 
be covered except for doors in the end adjoining buildings. 

Shipping Platform, in front of Converter Building, is cov- 
ered with a roof 240x31 ft. Trusses supported on 14-ft. posts 
12 ft. (7, and attached to Converter Building posts. Sides are 
left open 8 ft. from ground. 

All buildings to be iron constructions, well braced, and 
executed in a good and workmanlike manner. All roofs to 
be covered with 2|- in. corrugated iron No. 20, and sides, as 
specified, with 5^ iuo corrugated iron No. 22. All to be painted 
with good mineral paint. 



specifications of machinery for new copper converter 
plant for anaconda mining company, at 
anaconda, montana. 
To consist of 

36 Converters. 

12 Sets of stands and trunnions for same. 
12 Hydraulic cylinders, rack, gears and valves. 
6 Copper cupolas, and 3 extra wells. 
3 Blast furnaces, 6 forehearths. 
12 Double mould cars. 
12 Sets of runners. 



104 LEAD AND COPPER SMELTING. 

12 Steel-plate flues from converters to main flue. 

2 Sheet-steel blast pipes. 

2 12-ton traveling cranes, 

3 No. 7 Koots blowers. 
2 Chilian mills. 

12 Matte carts. 

100 Slag pots. 

6 Colorado Iron Works dumping slag-trucks. 



MACHINERY. 

Plate II shows the general plan and location of machinery 
in buildings. 

Converters : The converters are to be constructed as per 
Plates III, ly, y, and yi ; to be made in two sections. 
The lower section 6 ft. diameter by 5 ft. high ; to be made of 
7-16 tank steel. To it is riveted a dished and flanged head, 
also made of 7-16 steel. In the center of this head is a 1-in. 
hole for drainage of moisture from lining. The nose or upper 
section to be made of |-in. tank steel. Half way around the 
24-in. pouring opening is riveted a § x 12 in. wide reinforce- 
ment plate. Both sections are reinforced by 3 x 3 x |-in. 
angle iron rings, securely riveted to them. A band 7-16 x 10 
in. wide is riveted to lower section where lugs are attached. 

Lugs : Eight wrought iron or forged steel lugs to be riveted 
to each section located respectively as per drawings. Four 
of the lugs on nose section to have a loop or eye made in their 
upper ends for attaching chain hooks for lifting converters. 
Another eye for lifting the nose section alone is riveted on 
top, and should be so located that this section will hang level 
when lifted. 

Guides : To the nose section are further riveted four straps 
4 x|-in. iron, projecting 2 in., overlapping the lower section, 
so as to guide and hold sections in central positions when put 
together. All rivets used in construction of shells to be 
11-16 in.. Burden brand, 3 in. pitch. 

Lining Anchors to be made of 3-16x2-in. steel (see Plate 
III) and riveted to the interior of nose section and bottom 
of lower section, spaced about 10 in. C^ 9-16-in. rivets. 



SPECIFICATIONS. 105 



Tuyere Boxes : Eight tuj^ere boxes to be riveted to shell. 
Dimensions and location as per Plate VI. Eight holes, corre- 
sponding to tuyere holes in boxes, to be cut in converter 
shells, and the edges of said holes to be calked and made air- 
tight. 

Tuyeres are to be made of cast iron, as per drawing. 

TuYERE-Box Covers to be fitted with two valves, as shown 
in details, Plate YI, for facilitating the punching of tuyeres. 

Bustle Pipe : A rectangular cast-iron box 4x7 in. inside 
dimensions, to be made in three sections, provided with eight 
short nozzles corresponding to receptacles in tuyere boxes, 
and secured with two f -in. bolts through each box, making an 
air-tight joint, with suitable packing. 

Trunxion Ring : Cast-iron trunnion rings made in two sec- 
tions ; joints to be faced and bolted together with eight l^-in. 
turned bolts, fitted in reamed holes. Trunnion tips to be 
turned to gauge and templets in order to be interchangeable. 
Trunnion rings are provided with eight sleeve holes for 2-in. 
key bolts; bolts to be made of Norway iron, and keys of 
steel. The bolt holes in above-mentioned lugs to correspond 
exactly with sleeve holes in trunnion rings. 

Blast Connection between trunnion ring and tuyere box 
to be made with 5-in. pipe and fittings. 

Journals to be turned, bored, and planed as indicated on 
drawing, and made to fit trunnion ring and tips perfectly. A 
wrought-iron strap to be fitted to box part of trunnion where 
jib and key pass through, as shown in detail drawing. 

Boxes to be bored to fit trunnion iron to iron and planed in 
joint and underneath. 

Stands (Plates YII and YIII) to be planed on top for boxes 
and also on projection or washers for reach rods for hydraulic 
cylinder where indicated on drawing. 

Gear : Cast-iron gear, neat fit on trunnion, and secured by 
two keys ; shrouds to be turned ; diameter to be equal to 
pitch-line. 

Note: Referring to Plate II, general plan of machinery, it 
will be seen that it will be necessary to make portions of 
stands and other parts right and left. 

Hydraulic Cylinder : Plate IX is a detail drawing of 
liydraulic cylinder, rack and gear, etc. Cylinder, 15-in. bore, 



106 LEAD AND COPPER SMELTING. 

7 ft. i in. iong (6 ft. 4i-in. stroke) 1^ in. thick; cylinder 
to be bored true, and piston to be fitted with two brass rings 
bored and turned concentric, rubber or square fiber x^acking 
to be used between piston and rings ; ground joint to be made 
between follower and piston ; follower to have two |-in. tap- 
holes on top for eye bolts. Cylinder heads to have spigot 
joints and be bolted to cylinder with twelve 1-in. turned bolts. 
Piston rod, 4-in. steel, 9 ft. 6^ in. long. Cylinder to be counter- 
bored, so piston will overrun and leave no shoulders. 

Rack : 9 ft. 1 in. long, to be planed on shrouds and back 
where roller rests attached to piston rod by socket and key. 
A keyhole is also made above rod, in order to back out rod 
when necessary. Cylinder is attached to stands by four 2^-in. 
rods. Four pieces of extra strong 3-in. wrought-iron pipes 
act as sleeves ; these must be faced off exactly to one length. 
Hydraulic Yalve : Plate X is a detail drawing for a 4-way 
valve. Slide valve, stufiing box, nut and gland to be brass 
castings ; valve stem steel forging ; the teeth on it to be cut 
out of the solid. 

Runners : Plate XI is a detail drawing for runners and 
their supports ; also platform from cupola foundation to 
runners. One straight runner 10 x 12 in. by 12 ft. 8 in. long, 
made of 3-16 tank steel and 2 x 2-in. angle irons attached to 
cupola with a cast-iron socket leads to the double runners, 
which also are made of 3-16-in. tank steel and 2x2 angle 
irons with curved cast-iron mouthpieces. As shown, they are 
pivoted on a casting suspended between two 8-in. channel 
irons. At the extremity of these channel irons are attached 
hangers for a circular track, on which the supporting rollers 
rest. 

Converter Flues : Plate XII is a detail drawing of the 
steel-plate flue leading from converters to main flue. It 
is made of 3-16-in. tank steel and 2x2 angle irons, all but the 
first two feet nearest converter, which is made of 5-16 tank 
steel. Flue to be suspended from cupola floor beams, and 
traveling crane girder by suitable straps. 

Traveling Cranes : Two 12-ton cranes 42 ft. 3 in. span, 
required for lifting converters out of and into their bearings. 
Cranes to be operated by electric motors. Two speeds re- 
quired for hoisting ; one for lifting the whole converter, and 



SPECIFICATION. 107 

a faster speed for handling the nose section alone. Trolley 
to be traversed, and cranes to travel the whole length of 
building. Plate XIII is a cross-section of building, side 
elevation of crane, and of one girder for traveling crane and 
position of posts. Stay braces for supporting this girder to 
be bolted to retaining wall in front of cupolas ; the other 
girder, as mentioned above, to be supported on bracket 
riveted to building posts. 

Mould Cars, constructed as per drawing of 6-in. channel 
beams and 3-16-in. tank steel, cast-iron axle boxes (babbitted) 
and flange wheels. Twelve double cars required. 

Floor Plates : Corrugated floor plates, manhole plates, 
etc., as per drawing. 

Cupolas : Plate XIY is a detail of cupola ; six cupolas 
required. The well, 5 ft. 4 in. diameter by 4 ft. 6 in. high, 
is made of 5-16 tank steel, in halves, jointed on the sides by 
S X 3 angle irons, riveted to shell and bolted together. 

Water-Jacket is made tapered, inside diameters being 46 
in. and 54 in. ; total height 3 ft. 3 in. ; to have a water space 
4i in. Inside shell made of |-in. flange steel, and outside 
shell 5-16-in. steel, provided with four cast-iron ears for feed 
water and overflow, located as per drawing ; also four hand 
holes, plates, and crabs ; 11-16-in. Burden rivets. If in. pitch, 
to be used. 

Cupola Shell : Q6 in. diameter by 6 ft. high, made of 5-16- 
in. tank steel. Pour cast-iron brackets are riveted to it and 
eight wrought-iron knees ; similar knees to be attached to 
water jacket for hanging same to cupola shell by eight 1-in. 
rods. To the lower end inside and to the upper end outside 
of shell are riveted 3 x 3-in. angle iron rings. 

Stack : A 36 in. diameter by 24-ft. high stack, made of No. 
8 steel, is connected to shell by a 3 ft. long cone, its lower end 
being flanged and bolted to the above-mentioned angle iron 
ring. Top of stacks to be provided with cast-iron dampers, 
suitable levers and f ulcrums ; a band | x 3 in. is riveted to top 
of stacks. A branch 36 in. diameter leading to main flue 
made of 3-16 tank steel, provided with a butterfly valve, as 
shown on Plate XIII. 

Charging-door opening in cupola is reinforced by a ^ x 2-in. 
wrought-iron frame, riveted to inside of shell. 



108 LEAD AND COPPER SMELTING. 

Tuyeres : Four tuyere points with peep holes for each 
cupola. Tuyere points are connected to bustle-pipe nozzles 
by canvas sacks. 

Charging Floor (Plate XY) consists of 5-16-in. corrugated 
wrought-iron plates on 8-in. steel I-beams supported on 
wrought-iron columns, provided with cast-iron top and bottom 
caps. Columns supporting cupola to be cast iron, with brack- 
ets for floor beams, as shown on plan. 

Blast Furnaces : Plate XYI shows details of furnaces. 
Three furnaces complete. 

Crucible : Crucible plates to be made of cast iron li in. 
thick, well ribbed and bound together by bolts ; they fit in a 
pan made of 3-16-in. tank steel, with 3 x 3-in. angle iron frame, 
riveted to its edges ; one end to be provided with a cast-iron 
spout or taphole. 

Water-Jacket : To be made in four pieces. Sides to be 
made of f-in. Park Bros.' flange steel ; each side to have six 
tuyeres 4x5 in. rolled and beaded. Inside sheet of end 
pieces to be |-in. Park Bros.' flange steel ; outside sheet 5-16- 
in. tank steel, riveted together on a 2 x 2-in. angle-iron frame. 
,7ater-jacket to be 5J in. thick, and each side piece to be pro- 
vided with two cast-iron ears for feed and overflow, and each 
end piece to be provided with one cast-iron ear. Each piece 
of water-jacket to have li-in. gas-pipe washout holes and 
two hand holes, plates and crabs. One end piece to have an 
arch over cast-iron spout, as shown. 11-16-in. Burden rivets. 
If in. pitch, to be used. Jacket to be well stay-bolted and 
bound together with key rods, as shown. 

Brickwork : Above water-jacket will be built on a cast- 
iron plate supported on a rectangular frame, consisting of 
three 8-in. steel I-beams mitered together ; an angle iron 
knee to be riveted in each corner of inside frame. I-beam 
frame to rest on four cast-iron columns 9 in. diameter 7 in. 
core, having suitable top and bottom base plates. BrickAvork 
to be ]>ound together by 3 x 3 angle iron on corners and eye- 
bolt rods as shown. 

Door Sills : Cast-iron plates 22 x 48 x 2 in. thick for charg- 
ing doors, two for each furnace. 

Stack Plate : A cast-iron frame 8 ft. 4 in. by 6 ft. 2 in. 
by 1 in. thick, with a 1 J x |-in. rim all around outside edge 



SPECIFICATIONS. 109 

and a 2 X |-in. rim on top around a 6 ft. by 3 ft. 10 in. hole for 
the 6 ft. 2 in. by 1 ft. stack to fit into. 

Stacks to be 48 in. diameter by 20 ft. high, made of Xo. 8 
steel and provided with cast-iron dampers, levers, and ful- 
crums. A I X 3-in. band around top of stack. A 48 in. 
diameter branch, 3-16 in. steel, leading to main flue provided 
with a butterfly valve. 

Chargeng Floor : A 3-in. plank floor on 3 x 12-in. joists are 
carried on 10-in. steel I-beams supported on retaining wall, 
intermediary 6-in. wrought-iron columns and attached to 
building posts. A strip 4 ft. wide by 84 ft. long in front of 
chute from stockhouse to be covered with 3-16-in. wrought- 
iron plates, and a space 4 ft. wide by 18 ft. long on each side 
of furnaces, and 4 x 74 ft. in front of furnaces, to be covered 
with |-in. cast-iron plates, as indicated on drawing. 

FoREHEARTHS : 4 X 3 ft. 6 in. X 2 ft. high to be made in halves 
of 5-16-in. tank steel and 2J angle irons, jointed on the sides 
by 3 X 3-in. angle irons riveted to box and bolted together. 
Forehearths to be provided with two cast-iron detachable 
spouts, located as per drawing. 

Tuyere Points : Twelve tuyere points with peep holes for 
each furnace, connected to bustle-pipe nozzles by canvas 
sacks. 

Blast Pipes: Plate XVII general plan and elevation of 
blast pipes and details of bustle i)ipes for cupolas, and blast 
furnaces and converters. 

Xote: Connections to blowers and blowing engine receivers 
with above pipes to be made from actual measurements when 
in place. 

Material : Blast iDipes to be made of black sheet steel ; 
thickness for the different sizes of pipes are marked on 
drawings. 

Blowers : Two Xo. 7 Roots blowers, bottom discharge. 

Chilian Mills : Two Chilian mills, of T. Carlins & Sons' 
make (Allegheny, Pa.), wet grinding pan 7 ft. 6 in. diameter, 
as described in their catalogue, pag^e 115. 

Matte Carts : Twelve, as per Plate XYIII. 

Slag Pots : One hundred, as per drawing. 



110 LEAD AND COPPER SMELTING. 

INTENTION OF THESE SPECIFICATIONS. 

(1) It is the intention of these specifications to have all 
converters, furnaces, and all the necessary appliances for a 
complete plant ; to be made of the best material and work- 
manship. All the castings to be made true to form, to be 
free from blowholes, cold laps, or scales, or imperfections of 
any kind. The forgings to be smooth, true to form, and free 
from defects of all kinds, and the best material must be used 
in same. 

(2) All the several parts of converters, furnaces, etc., 
must be made interchangeable, and all finished -parts to be 
made to templets. 

(3) When the wording of any clause of the specifications 
may not be clearly understood, or is open to misconstruction, 
or when the plans are at fault from errors or mistakes, the 
company is hereby authorized to give the necessary explana- 
tion through its engineer, his decision to be final in each case ; 
and all the work contemplated and described by plans and 
specifications shall be done to the satisfaction of the Ana- 
conda Mining Co., or its engineer, who shall judge as to the 
fitness of materials used in the construction of their several 
parts, and he shall have the right of correcting any errors or 
omissions in the plans or specifications, when such corrections 
are necessary for the proper fulfillment of their intentions. 

(4) Further detailed drawings, if necessary, will be made 
by the contractors, and approved by the company's engineer, 
or be furnished by him, as the case may require, in order to 
carry out the work and intent of the specifications. 

(5) These specifications are accompanied by blue prints, 
marked "Copper Converter Plant for Anaconda Mining Co." 

(6) All work must be set up in shop with necessary 
attachments, and properly marked so as to facilitate erection. 



COPPER CONVERTER PUNT 

FOR 
ANACONDA MINING CO. 

Jan. 14, 1893. 



PLATE I 




PLATE ] 




GENERAL PLAN 

OF 

COPPER CONVERTER PLANT 

FOR 

ANACONDA MINING CO. 
OCT. 15, n 






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GENERAL PLAN 
COPPER CONVERTER PLANT 





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DETAILS OF ANCHORS 
FULL SIZE 





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PLATE V 




DETAILS OF TRUNION RING ETC. FOR CONVERTERS 
COPPER CONVERTER PLANT 

FOR 

ANACONDA MINING CO. 

Mar. 29, 1893. 



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PLATE VI 




DETAILS OF TUYERBOXES AND TUYERS FOR CONVERTERS 
COPPER CONVERTER PLANT 

FOR 

ANACONDA MINING CO. 

Feb. 13,1893. 



2R3Tfl3VHOO ROl 2fl3YUT QUA 83XOafl3YU1 
- AaMOOAHA 




DETAILS OF COPPER CONVERTER-STAND, ETC, 
COPPER CONVERTER PLANT 

ANACONDA MININQ 



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TAILS OF HYDRAULIC CYLINDER RACK AND GEAR 

FOR 

CONVERTERS 
COPPER CONVERTER PLANT 

FOR 

ANACONDA MINING CO. 

Oct. 24, 1892. 



;k gear 

PITCH=a 
ETH=29 NO, OF TEETH=34- 

DIA. PITCH CIR.=32. 46' | 




DETAILS OF HYDRAULIC CYLINDER RACK AND C 





Lie FOUR-WAY VALVE 

:r converter plant 

FOR 

CONDA MSNINC CO. 





ANACONDA MINING CO. 




ANACONDA MINING 



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PLATE XIV 



DETAILS OF COPPER CUPOLA 

COPPER CONVERTER PLANT 




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WATER JACKET FURNACE 




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PAGE 

57 

21 

jrters . . 94 

85, 87 

76 

53 

23 

56 

96 

31 

81 

32 

47 

58 

100 

issing 100 

94 

3r Slag 23 

100 

89 

104 

ant at 103 

ace- 108 

31 

9 

o. 35 

98 

97 

1 

5, 18, 42, 52, 56, 57 

90 

13 

15 

9 



112 INDEX. 

PAGE 

Bullion and Slag, Relation between Assays at Leadville „ 20 

Calculation of Furnace "'barges , 18 

Casting Anodes direct "roui Converters .... 97 

Cast-Iron Moulds for Pouring Converted Copper 92 

Cast-Iron Water- Jackets 13 

Cbarges for b urnaces, Calculation of 18 

Charges for Lead Furnaces, Composition of , . 54 

Cbarging a Furnace, Metbod of 7, 22, 52 

Clay used for Converter Lining at Anaconda, Mont., Composition of 71 

Coke, Effects of Different Kinds on Furnace Running 58 

Coke, Percentage required in Lead Smelting 58 

Coke from Sabinas, Mexico 58 

Colorado Smelting Company, Pueblo, Colo 51 

Coloration of Copper Converter Flames 74 

Composition of Clay used for Converter Lining at Anaconda, Mont, .... 71 

Composition of Material for Converter Lining 71 

Converter Copper Bars, Sampling of 83 

Converter Copper, Grade in Gold and Silver <, 93 

Converter Copper, Kinds of finisb on 81 

Converter Linings, Life of 95 

Converter Linings, Life at Aguas Calientes, Mexico 96 

Converter Linings, Life at Great Falls, Mont 96 

Converter Lining, Life at Parrot Works, Mont 96 

Converter Plants, Design of 84 

Converter Slag, Resmelting of 26 

Converter, Water Jacketed, Experiment witb 67 

Converters, Copper, at Aguas Caliente.s, Mexico 85 

Converters, Copper, Coloration of Flames from. 74 

Converters, Copper, Design of, at Anaconda, Mont 104 

Converters, Copper, at Great Falls, Mont 86 

Converters, Copper, Handling by Hydraulic Pistons 90 

Converters, Copper, Objection to very large size 88 

Converters. Copper, at Parrot Works, Mont 86 

Converters, Copper, Slow Running and Remedy 76 

Converters, Method of Lining 94 

Converting Copper Matte at Anaconda 66 

Converting Copper Matte, Itemized Cost of 73 

Cope and Core Apparatus for Making Copper Moulds 91 

Copper Bars, Converter, Sampling of , . . 83 

Copper Bars from Converters, Distribution of Silver in 82 

Copper Converters at Aguas Calientes, Mexico 85 

Copper Converters, Coloration of Flames from 74 

Copper Converters, Design of, Anaconda, ^. ont 104 

Copper Converters at Great Falls, Mont 86 

Copper Converters, Handling by Hydraulic Pistons. 90 

Copper Converters, Kinds of finisb on Ni 

Copper Converters, Objection to very large size 8^ 

Copper Converters at Parrot Works, Mont 86 

Copper Converters, Pouring Slag from 75 

Copper Converters, Size and Production of 72 

Copper Converters, Slow Running and Remedy 76 

Copper Converting at Aguas Calientes, Mexico 76 

Copper Converting at Anaconda 66 

Copper Converting, Development of Heat in a Cold Charge 79 

Co(iper Converting, Grade of Matte Suitable for 80 

Copper Converting, Itemized Cost of 73 

Copper Converting, Loss in, at Anaconda. Mont 99 

Copper Converting, Metallurgical Reactions in 67 

Copper Converting Plant at Anaconda. Mont., Design of 102 

Copper Converting Plant. Arrangement of 90 

Copper Matte Smelting Furnaces, Jackets for 31 



113 



PAGE 
94 

91 
80 
1 
91 
93 
75 
99 
23 
91 
70 
73 

100 

100 
99 
25 
39 
81 
42 

100 
45 
52 
13 
15 

2, 5 
10 
21 
58 
57 
65 
18 
22 
35 



31 
25 
50 

47 

47 

4 

2 

52 

9 24 



Hi Iis"DEX. 

PAGE 

Hunt & Douglas' Process .,,.,,.... 15 

Hydraulic Piston for Handling Copper Converters. 90 

Jackets for Copper Furnaces 31 

Jackets, Design of , 34 

Jackets for Smelting Furnaces, Amount of water required 36 

Jackets for Smelting Furnaces at Leadville 13 

Lead Blast Furnaces, Design of 47 

Lead, Percentage required in Lead Smelting 6 

Lead Slag, Composition, at Leadville, Colo 4 

Lead Slags 54 

Lead Smelting, Losses in 54, 56 

Lead Smelting, Composition of Charges for 54 

Lead Smelting Furnaces, Obstruction in Crucibles. .. ....... 39 

Lining, Basic in Copper Converter, Experiment with 69 

Lining a Converter 94 

Lining, Converter, Composition of Material for 71 

Lining, Converter, Corrosion by Low Grade Matte 70 

Lining Converters, Metliod of 70 

Lining, Converter, Preparation of Material for 71 

Lining in Conper Converters, Life of 95 

Lining in Coppar Converters, Life at Aguas Calientes 96 

Lining in Copper Converters, Life at Great Falls, Mont 96 

Lining in Copper Converters, Life at Parrot Works, Mont 96 

Lining, Silicious, Part played in Copper Converting 69 

Loss in Copper Converting at Anaconda, Mont 99 

Loss in Copper Refining at Anaconda. Mont 99 

Loss in Ore Dressing at Anaconda, Mont 99 

Loss of Silver in Copper Converting 73 

Loss in Smelting at Anaconda, Mont 99 

Losses in Lead Smelting .54, 56 

Machinery in Copper Converting Plant at Anaconda, Mont. 103 

Mahala Mine, Leadville, Colo 2 

Maid of Erin Mine, Leadville, Colo 2 

Matte, Copper, grade suitable for converting 80 

Matte, Extraction of Gold and Silver from 15 

Matte Fall, Percentage required in Matte Smelting 6 

Matte Pots used at Omaha & Grant Works, Denver 43 

Matte, Grade required for Copper Converting 66 

Matte, Low Grade, Corrosion of Converter Lining by 70 

Matte Pots used at Arkansas Valley Works, Leadville 42 

Matte, Resmelting with Hot Blast 65 

Matte Settling from Slag, Principle of 24 

Matte and Slag, Relation between assays of , 20 

Matte and Slag, Relation between in Copper Converting 70 

Matte Smelting, Furnaces for, at Leadville 2 

Matte Smelting Furnaces at San Luis Potosi, Mexico 26 

Matte Smelting at Leadville, Amount of Fuel required 13 

Matte Smelting at Leadville, Capacity of Furnaces 13 

Matte Tapping from Furnaces, Method of 25 

Matte-tap, Position in Leadville Furnaces 26 

Matte, Tenor in Gold and Silver 20 

Mechanical Roasting Furnaces 59 

Mexican Metallurgical Company, San Luis Potosi, Mexico 43. 45 

Mining Anaconda Ore, Cost of 100 

Monterey, Mexico, Lead Smelting at 58 

Moulds for Casting Copper. . . 91 

Moulds, Cast-iron for Pouring Converted Copper 92 

Moulds, Copper, for Pouring Converted Copper 93 

Moulds, Steel, for Pouring Converted Copper. 93 

Pacbuca. Mexico, Ore used in lining Converters at Aguas Calientes 81 

Parrot Works, Mont., Copper Converters at 84, 86 



115 

PAGE 

96 

16 

71 

43, 45 

7 

59 

42 

99 

8, 50 

65 

81 

62 

59 

8 

39 

26 

28 

13 

31 

31 

5 31 

24, 25 

24 

69 

82 

76 

42 

20 

tion of 94 

>f 94 

26 

if 23 

4, 55 

20 

70 

75 

22 

[l. '.[[['.['[...'. 29 

28 

45, 46 

57 

54 

22 



116 IKDEX. 

PAGE 

Tuyere Bags » , » o . _ _ « 36 

Tuyeres, Kinds for Smelting Furnaces , , . , 36 

Tuyeres, Number required in a Smelting Furnace 32 

Types of Furnaces. 24 

Velardena, Mexico, Lead Smelting at , 58 

Volatilization of Lead in tlie Blast Furnace, Experience at Aguas Calientes 48 

Water, Amount required for Granulating Slag 46 

Water, Amount required for Jackets 36 

Water, Amount required for Sluicing Away Slag 46 

Water Coil Protection for Brick Furnaces 35 

Water Jackets 13 

Water- Jackets, Design of 34 

Water Jacketed Converter, Experiment wltli 67 

Water Jacketed Furnaces at Anaconda, Mont 108 

Wind required in Smelting 8 

Wolftone Mine, Leadville, Colo. 2 



Iting. 



?S, Jr. 

eatly Enlarged* 
world on 



1 directions how to 
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i'uel.) 



Scope; Apparatus; anc? 



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